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CXI 
BUREAU OF MINES 3 

INFORMATION CIRCULAR/ 1989 




Surface Testing and Evaluation 
of the Multiple-Unit Continuous 
Haulage System 



By Jasinder S. Jaspal, Lee A. Erhard, 
and William D. Mayercheck 



UNITED STATES DEPARTMENT OF THE INTERIOR 



Mission: As the Nation's principal conservation 
agency, the Department of the Interior has respon- 
sibility for most of our nationally-owned public 
lands and natural and cultural resources. This 
includes fostering wise use of our land and water 
resources, protecting our fish and wildlife, pre- 
serving the environmental and cultural values of 
our national parks and historical places, and pro- 
viding for the enjoyment of life through outdoor 
recreation. The Department assesses our energy 
and mineral resources and works to assure that 
their development is in the best interests of all 
our people. The Department also promotes the 
goals of the Take Pride in America campaign by 
encouraging stewardship and citizen responsibil- 
ity for the public lands and promoting citizen par- 
ticipation in their care. The Department also has 
a major responsibility for American Indian reser- 
vation communities and for people who live in 
Island Territories under U.S. Administration. 



Information Circular 9228 

w 



Surface Testing and Evaluation 
of the Multiple-Unit Continuous 
Haulage System 



By Jasinder S. Jaspal, Lee A. Erhard, 
and William D. Mayercheck 



UNITED STATES DEPARTMENT OF THE INTERIOR 
Manuel Lujan, Jr., Secretary 

BUREAU OF MINES 
T S Ary, Director 



i 



Library of Congress Cataloging in Publication Data: 



Jaspal, Jasinder S. 

Surface testing and evaluation of the multiple-unit continuous haulage system. 

(Bureau of Mines information circular; 9228) 

Includes bibliographical references. 

Supt. of Docs, no.: I 28.27:9228. 

1. Mine haulage. I. Erhard, Lee A. II. Mayercheck, William D. III. Title. 
IV. Title: Multiple-unit continuous haulage system. V. Series: Information 
circular (United States. Bureau of Mines); 9228 

- TN295.U4 - [TN331] 622 s [622'.66] 89-600120 



CONTENTS 

Page 

Abstract 1 

Introduction 2 

Acknowledgments 2 

Description 2 

Lead vehicle 3 

Intermediate vehicles 4 

Discharge vehicle 5 

Bridge conveyor 5 

Surface testing 6 

Jeffrey mining machinery testing 6 

METF test program overview 6 

Tram and maneuverability trials 6 

Steering system 7 

Tracking-retracking tests 10 

Simulated production cycle trial 13 

Conveyor system tests 16 

Conveyor noise-level survey 16 

Conveyor speed test 18 

Conveyor time sequencing 20 

Coal conveying tests 20 

March 14 to April 12, 1984 20 

July 31 to August 3, 1984 28 

September 10-11, 1984 31 

Umbrella miner cutting trials support 31 

Conveyor system problem summary 33 

Coal conveyor acceptance test-October 29, 1986 34 

Drawbar pull test 37 

Rough terrain test 38 

Canopy load test 39 

Ground pressure evaluation 39 

Surface test summary 40 

System modifications 41 

MSHA experimental permit approval 41 

Operator canopy redesign 42 

Elevation of conveyor deck 42 

Cable handling tray 43 

Discharge vehicle steering system 43 

Emergency shutdown system 43 

Conveyor chain slack adjustment mechanism 47 

Conveyor chain holddowns 47 

Conveyor deck extension 50 

Conclusions 50 

Appendix A.-MUCH system specifications 51 

Appendix B.-Conveyor test breakdown and repair log 52 

Appendix C.-MUCH system coal carryback loss-conveyor test-July 31 to August 3, 1984 56 

Appendix D.-Conveyor acceptance test log-October 29, 1986 57 

Appendix E.-Recommended spare parts for in-mine trial 58 

Appendix F.-MUCH system modifications summary 59 



ILLUSTRATIONS 

Page 

1. Multiple-unit continuous haulage system 3 

2. Lead vehicle 3 

3. Intermediate vehicle , 4 

4. Discharge vehicle 6 

5. Bridge conveyor 6 

6. MUCH system operational maneuvers in equipment maneuverability trial area 8 

7. Mechanical steering linkage geometry 9 

8. Tracking layout, 12- and 43-ft turning radii 10 

9. Tracking capability, 43-ft turning radius 11 

10. Trial A and trial B inby 12 

11. Trial A and trial B outby 14 

12. Lead vehicle behind shuttle car 15 

13. Tracking variance 16 

14. Conveyor noise-level test configuration 17 

15. Conveyor test configuration 21 

16. Test configuration and instrumentation system 21 

17. Typical strip-chart data 25 

18. Coal fines buildup on conveyor deck 26 

19. Typical coal carryback 27 

20. Conveyor drive gearbox temperature rise 28 

21. Conveyor return coal buildup 29 

22. Coal fines buildup in pelican beak 29 

23. Cleanout ports in pelican beak 30 

24. Coal carryback loss after haulage testing, July 31 to August 3, 1984 31 

25. Water spray nozzles 32 

26. Water spray manifold in pelican beak 32 

27. Conveyor chain riding over rock 34 

28. Rock caught between shaft and flight 35 

29. Gap between deck and shaft 35 

30. Conveyor deck plate 36 

31. Conveyor chain holddowns 36 

32. Crane and dynamometer used during drawbar pull test 38 

33. Drawbar pull test 38 

34. Lead vehicle operator canopy load test 40 

35. Original operator canopy 43 

36. Lead vehicle operator canopy 44 

37. Operator in lead vehicle 44 

38. Clearance between hopper and deck-original design and modified design 45 

39. Conveyor deck before and after modification 45 

40. Height of throat area-original design and modified design 45 

41. Cable handling tray 46 

42. Discharge vehicle hydraulic steering system arrangement 46 

43. Emergency shutdown pull cord 47 

44. Conveyor chain slack adjustment screw 48 

45. Conveyor chain slack adjustment mechanism 48 

46. Chain holddown mounting hole 49 

47. Conveyor deck extension plate 49 



Ill 

TABLES 

Page 

1. MUCH tracking-retracking trials 15 

2. Conveyor noise test results 17 

3. Conveyor speeds 19 

4. Particle transport speed 20 

5. Conveyor startup time 20 

6. Haulage test summary 23 

7. MUCH haulage testing data summary-electrical parameters 24 

8. Conveyor carryback loss 26 

9. Tire ground pressure for loaded and unloaded vehicles 40 






UNIT OF MEASURE ABBREVIATIONS USED IN THIS REPORT 


A 


ampere 


kW-h/st 


kilowatt hour per short ton 


°C 


degree Celsius 


kW-h/(st-ft) 


kilowatt hour per short ton per foot 


dBA 


decibel, A-weighted 


lb 


pound 


op 


degree Fahrenheit 


min 


minute 


ft 


foot 


mm/s 


millimeter per second 


ft/min 


foot per minute 


pet 


percent 


gal/min 


gallon per minute 


psi 


pound per square inch 


h 


hour 


r/min 


revolution per minute 


hp 


horsepower 


s 


second 


Hz 


hertz 


st/min 


short ton per minute 


in 


inch 


V 


volt 


in 2 


square inch 


Vac 


volt, alternating current 


kVA 


kilovolt ampere 


yd 3 


cubic yard 


kW 


kilowatt 







SURFACE TESTING AND EVALUATION OF THE MULTIPLE-UNIT 
CONTINUOUS HAULAGE SYSTEM 



By Jasinder S. Jaspal, 1 Lee A. Erhard, 2 and William D. Mayercheck 3 



ABSTRACT 

Most of the underground coal in the United States is mined via room-and-pillar mining methods with 
continuous miners. These machines operate intermittently because they have to wait for shuttle cars to 
interchange positions. To overcome this discontinuity in shuttle car haulage and to realize the full 
production potential of continuous miners, the U.S. Bureau of Mines developed a multiple-unit 
continuous haulage (MUCH) system through a research contract with Jeffrey Mining Machinery Div. 
(JMMD). The MUCH system consists of 12 rubber-tired vehicles and a bridge conveyor. The rubber- 
tired vehicles are connected by a unique mechanical linkage system to form a 250-ft train. The 
mechanical linkage permits the vehicles to track-retrack the preceding vehicle in both inby and outby 
directions. The cut coal cascades from one vehicle to another until it is discharged on the section 
conveyor belt. The MUCH system provides continuous haulage to the continuous miner. The MUCH 
system was initially surface tested by JMMD. Subsequent extensive testing was conducted by the Bureau 
at its test facilities at Bruceton, PA. Deficiencies found during testing were corrected. This report 
summarizes initial tests by JMMD and subsequent extensive tests and evaluation by the Bureau. 



x Mining engineer, U.S. Bureau of Mines, Pittsburgh Research Center, Pittsburgh, PA. 
2 Manager, Plant Services, Mining Test Facility, Boeing Services International. 
Supervisory physical scientist, Pittsburgh Research Center. 



INTRODUCTION 



In room-and-pillar mining, a continuous miner cannot 
operate continuously because it must halt production- 
loading operations to permit a loaded shuttle car to move 
out and an empty shuttle car to move in before the 
production-loading operation can be resumed. This 
intermission precludes realization of full production 
potential and shows a weakness in a shuttle car haulage 
system. 

For the mining of 90° or other angle pillar entries or 
crosscuts, no haulage system was available that could 
remove the production coal continuously from a continu- 
ous miner. This situation led the Bureau to develop the 
Multiple-Unit Continuous Haulage (MUCH) system 
through a contract with Jeffrey Mining Machinery Div. 
(JMMD) of Dresser Industries, Inc. JMMD designed and 
fabricated the system and surface tested it at its plant. 
During these tests, numerous modifications were made to 
the MUCH tracking-retracking and conveying systems. 
After these modifications, the MUCH system was tested 
at the Ohio Transportation Research Center (OTRC). 
Additional modifications were made to the system to im- 
prove its performance. 4 The modified MUCH system was 
brought to the Bureau's Mining Equipment Test Facility 
(METF) in 1983 for more comprehensive surface testing 
with the purpose to evaluate the system performance 



and operational characteristics, correct any faults found 
during tests, and prepare the system for an in-mine trial, 
including an experimental permit approval from the Mine 
Safety and Health Administration (MSHA). 

Most mine operators prefer not to risk the installation 
and operation of prototype equipment because any failures 
would be expensive in terms of operating costs and loss of 
production. Because of this risk, there is a strong ten- 
dency in the mining industry to resist the introduction of 
new equipment or reject promising equipment after short 
duration of unsuccessful trials. The MUCH system was, 
therefore, subjected to rigorous surface tests to improve 
its reliability in performance and eliminate or reduce any 
major design deficiencies. Modifications and surface tests 
on the MUCH system were completed in December 
1986 at the METF. The results of tests indicate that the 
MUCH system will perform successfully in an in-mine 
trial. 

The MUCH system is scheduled to be used in 1989 at 
a highwall operation with a remotely controlled Jeffrey 
102HP continuous miner. Necessary changes are being 
made on it for its operation with a remote controlled 
system. 



ACKNOWLEDGMENTS 



The authors want to extend their sincere appreciation 
to Robert J. Evans, civil engineer, Pittsburgh Research 



Center, for technical guidance during surface testing at 
the METF. 



DESCRIPTION 



The MUCH system consists of 12 rubber-tired vehicles 
in which there is one lead vehicle, 10 intermediate vehicles, 
and one discharge vehicle with a bridge conveyor to form 
a 250-ft-long train. Intermediate vehicles can be added or 
removed from the train to suit the section requirement. 
Each MUCH vehicle has a chain conveyor mounted on a 
transporting vehicle that features four-wheel steering and 
two-wheel drive. There is a hopper on each vehicle to 
receive coal from an inby vehicle. Except for the lead 
vehicle, the hopper of each vehicle is sized and shaped to 
accept coal in all steering positions. 



The vehicles are connected by a patented 5 self-tracking 
steering system that connects adjacent vehicles into a 
train with automatic mechanical tracking and retracking 
(fig. 1). Coal cascades from conveyor to conveyor at the 
rate of 12 st/min until it reaches the section belt. The 
train of vehicles is steered inby by the MUCH system 
operator in the lead vehicle, and outby by the discharge 
vehicle operator located outby near the section belt. The 
mechanical steering linkages on each vehicle enable all 
vehicles to sequentially track the path of the preceding 
vehicle through a mine at 80 ft/min. System specifications 
are given in appendix A. 



4 Hundman, G. J., and P. W. Meisel. Development of a Multiple 
Unit Continuous Haulage System (contract J0333941, Dresser Indus- 
tries, Inc.). BuMines OFR 101-84, 1983, 326 pp; NTIS PB 84-188630. 



Voight, E. T. Steering System for a Train of Rail-less Vehicles. 
U.S. Pat. 4,382,607, May 10, 1983. 



Section belt or chain structure 



Crosscut 



Lead vehicle 
operator 




Continuous miner 



Discharge vehicle operator 



Mine entry 



Figure 1 .-Multiple-unit continuous haulage system. 



LEAD VEHICLE 

The lead vehicle is the first vehicle in the train and it 
receives coal directly from the tailboom of a continuous 
miner. An operator compartment with canopy is 
positioned on the right side of the vehicle. The lead 
vehicle contains a hydraulic power system to power the 
hydraulic front-wheel steering and the hopper lift cylinders. 
The lead vehicle also has headlights, a communications 
system, a panic bar emergency shutoff system, and 
hydraulic brakes (fig. 2). 

The MUCH system is controlled by an operator from 
the lead vehicle. The operator sits on a padded seat, the 
back of which adjusts to a reclining position for operating 
in lower seams. 

A tram control stick is located to the left of the 
operator, near the left knee. Movement of the stick 
forward provides forward tram, while movement to the 
rear causes reverse tram. The right foot pedal is a dead- 
man's control that commands tram movement of the train. 
Simultaneous deflection of the control stick and depression 
of the right foot pedal are required to tram. This 
simultaneous action cannot be accomplished from outside 
the operator compartment. 

Lateral movement of the left control stick operates a 
hydraulic valve located behind the operator's seat. This 
hydraulic valve commands lead vehicle front-wheel steering 
through a pair of push-pull hydraulic cylinders. A control 
lever is also located to the right of the operator. This 
lever controls a hydraulic valve to raise and lower the lead 
vehicle's coal receiving hopper. 




PLAN 




ELEVATION 



Figure 2.-Lead vehicle. 



Not to scale 



In addition to the right foot pedal for tram control, a 
left foot pedal operates a disk brake through a hydraulic 
master cylinder. This brake is on the lead vehicle rear 
wheels only, and serves as a braking assist to the tram 
motor brakes. An electrical control case is mounted to the 
left of the operator. The control panel functions, starting 
at the outby end, are as follows: 



Lead vehicle motor control breaker. 
Hydraulic pump and headlight breaker. 
Conveyor systems start switch. 
Conveyor systems stop switch. 
Lead vehicle conveyor reverse switch. 
Lead vehicle conveyor run switch. 
Start hydraulic pumps switch. 
Stop hydraulic pumps switch. 
Lead vehicle headlight on switch. 
Lead vehicle headlight off switch. 

An emergency stop bar, which activates an emergency 
stop switch, runs along the upper edge of the control 
panel. The emergency stop switch activates the main line 
circuit breaker (CB-1) shunt trip, causing power to be 
interrupted over the entire train. 

Two headlights are mounted next to the conveyor at the 
front of the lead vehicle. The headlights are operated 
from the lead vehicle operator station. During normal 
operation, the headlights would remain on. When the 
system is positioned behind the continuous miner, 
receiving and transporting coal, the MUCH operator can 
switch the headlights off if desired. 

Tram Drive 

The horsepower of the tram motor is greater on the 
lead vehicle than on the intermediate vehicles. The higher 
horsepower motor is required because of the additional 
weight of the lead vehicle. The lead vehicle carries the 
system operator, operator compartment, system controls, 
and a larger front-receiving hopper, and it must travel at 
the same speed as the intermediate vehicles. Therefore, 
the lead vehicle tram motor is 7.5 hp instead of 5 hp, as 
used on the intermediate vehicles. 

All tires used on the MUCH system are the same 
size-8.25 by 15. However, the additional weight of the 
lead vehicle requires higher tire inflation pressure. The 
lead vehicle tire pressure is 100 psi, while the intermediate 
vehicles have 75-psi foam-filled tires. 

Hydraulic System 

The lead vehicle utilizes hydraulic power for steering 
and to raise and lower the coal receiving hopper. The 
hydraulic package located on the left side of the lead 
vehicle includes a 1-hp electric motor, hydraulic pump, 
hydraulic reservoir, and function controls. The hopper of 
the lead vehicle is 18 in longer than the hoppers of the 
intermediate vehicles. This additional length gives the 
MUCH system a capability to remain stationary while the 
continuous miner goes through one cycle of sumping, 
cutting, and loading. 



Conveyor Drive 

The lead vehicle and all other vehicles of the MUCH 
system have a 15-hp, electric, flange-mounted, conveyor 
drive motor attached to a speed reducer. A chain drive 
with a 1.16:1 sprocket speed reduction is used from the 
speed reducer to the conveyor drive shaft. The conveyor 
consists of 2-in-pitch, double-strand chain, 506 in long, 
and 29.5 in wide. The majority of flights are on 10-in 
centers. The conveyor deck is 30 in wide and has 9-in 
sideboards. The lead vehicle conveyor is the last to start 
in the MUCH system train. It is, however, the first vehi- 
cle from the continuous miner that starts conveying coal 
outby. Because it is the last inby vehicle of the train, it 
is not provided with a speed switch. With the chain run- 
ning at 280 ft/min, a conveying capacity of 12 st/min is 
achieved. 

INTERMEDIATE VEHICLES 

The intermediate vehicles constitute the bulk of the 
system, and consist of conveyor deck plates, conveyor 
sideboards, receiving hoppers, axles, speed reducers, drive 
motors, electrical control boxes, and frames (fig. 3). 

Tram Drive 

Each individual intermediate vehicle has a 5-hp tram 
drive motor with a spring-actuated, electrically released 
brake. The tram motor is face mounted to a speed 
reducer, which is coupled to the front-driving axle via a 
drive shaft. If tram power to an intermediate vehicle is 
lost, the system can continue to operate by disconnecting 
the tram drive universal joint on the drive shaft until the 
problem can be corrected. 




t^ 



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ELEVATION 



Not to scale 



Figure 3. -Intermediate vehicle. 



The brake is part of the electric motor and is spring- 
actuated and electrically released. The brakes are auto- 
matically applied with any electrical power interruption. 
The brake is used for parking and emergency applications 
only, and is controlled by the main breaker switch, panic 
bars, or any planned or unplanned power interruption. 

There are two 15:1 speed reducers mounted on each 
vehicle in the system. The tram motor speed reducer is 
mounted on the lower frame under the conveyor; the 
conveyor drive speed reducer is mounted on the left side 
motor mount shelf. 

Conveyor Drive 

The chain conveyor drive system of the intermediate 
vehicles is similar to the lead vehicle. A speed switch 
is mounted to the tail shaft of each conveyor on each 
intermediate vehicle. The speed switches are designed 
to close when the conveyor speed reaches 85 pet of nor- 
mal operational speed. When the conveyor start switch 
is activated, only the bridge conveyor starts. When the 
bridge conveyor reaches 85 pet of normal speed, the 
speed switch closes and the electrical circuit logic supplies 
power to the discharge vehicle conveyor drive motor. 
When this conveyor reaches 85 pet of normal speed, its 
speed switch closes to activate the next inby conveyor 
drive motor. As this action continues, the conveyors of 
the train sequentially start from outby to inby until the 
lead vehicle conveyor is actuated. The conveyor drive 
switching logic, in conjunction with the speed switches, 
is designed to prevent coal from being dumped onto a 
stopped conveyor. 

DISCHARGE VEHICLE 

The discharge vehicle, patterned after the intermediate 
vehicles, differs in a number of respects. This vehicle is 
designed to support the forward half of the bridge convey- 
or, which is connected to the vehicle by a hitch. The 
bridge conveyor increases the load on the discharge vehicle 
rear axle and decreases the load on the front axle. There- 
fore, the discharge vehicle (fig. 4) is the only vehicle with 
rear-wheel drive. 

The electrical control box on the discharge vehicle is 
larger than those of the other vehicles because it contains 
the switching controls for both the discharge vehicle and 
bridge conveyor; the main electrical power enters the train 
at the discharge vehicle. The control case has a main line 
circuit breaker (CB-1) at the main power entry point. A 
480-V ac to 120-V ac stepdown transformer is mounted on 
the discharge vehicle and supplies 120-V ac control power 
to the system. 

The primary MUCH system operator is located at the 
lead vehicle while the secondary operator, or helper, is 
located at the discharge vehicle. A pager phone is pro- 
vided at both the lead and discharge vehicle operator 
stations, as well as at the midpoint of the system. 

The discharge vehicle operator controls all steering of 
the system when tramming in the outby direction. Because 



the discharge vehicle is connected to the panel belt by the 
bridge conveyor, steering is needed to keep the system 
from running into the mine rib or panel belt. The steering 
is controlled by a hydraulic unit that includes a 1-hp 
electric motor and a hydraulic reservoir. The hydraulic 
unit is mounted on the side of the conveyor sideboard. 
When applied by the operator, the brake operates on the 
rear axle and is used to provide tension in the train when 
tramming inby or when on a slope. 

Two headlights are mounted on the outby end of the 
discharge vehicle. The headlights assist operation of the 
system in the outby direction. 

The discharge vehicle operator has no operator's com- 
partment or platform. Therefore, for operator protection, 
rear-wheel fenders are installed over the rear wheels. 

An emergency stop switch is mounted on each side of 
the discharge vehicle. These emergency stop switches 
activate the CB-1 shunt trip coil, causing power to be 
interrupted over the entire train. 

Tram Drive 

The horsepower of the tram motor on the discharge 
vehicle is similar to the lead vehicle because of the 
additional weight of the bridge conveyor. The discharge 
vehicle tram motor is 7.5 hp instead of 5 hp, as used on 
the intermediate vehicles. 

BRIDGE CONVEYOR 

The bridge conveyor (fig. 5) mounts on a hitch on the 
outby end of the discharge vehicle, and transports mined 
material to the section panel belt. The outby end of the 
bridge conveyor rides on a bridge dolly that rides the panel 
belt frame. The bridge dolly is built to accommodate a 30- 
in-wide panel belt. The bridge conveyor has an active 
length of 22 ft. The bridge conveyor chain is powered by 
a 15-hp electric motor mounted on the bridge conveyor. 
A supply unit is positioned on the bridge conveyor for the 
discharge vehicle hydraulic steering power unit. 

The bridge dolly frame contains four wheels that ride, 
two on each side, on angle iron guides attached to the side 
of the panel belt frame. The dolly frame also contains a 
crescent ring, in which a matching collar is anchored and 
allowed to angularly rotate. The outby end of the bridge 
conveyor pins to the collar. The bridge dolly is designed 
to run on a special belt tailpiece built by Long-Airdox, 6 
which is described in Long-Airdox bulletin 5-973-1. 
Twenty-nine of the 9-ft sections are connected together to 
permit the bridge dolly to travel approximately 260 ft over 
the section panel belt. The 260-ft distance is required to 
permit penetration of the vehicle train through lateral and 
parallel entries during advances of the tailpiece. The tail- 
piece would consist of 28 increment sections and 1 tail 
section. 



Reference to specific products does not imply endorsement by the 
U.S. Bureau of Mines. 




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ELEVATION Not t0 scale 

Figure 4.-Discharge vehicle. 



ELEVATION 



Figure 5. -Bridge conveyor. 



Not to scale 



SURFACE TESTING 



JEFFREY MINING MACHINERY TESTING 

During the 1977-79 period, JMMD conducted a series 
of tests on the MUCH system at its facility. The objec- 
tives of these tests were to verify the tracking-retracking 
ability of the system in 90° room-and-pillar configuration 
in an aboveground environment and ascertain the convey- 
ing capabilities of the system. 

These tests revealed that each vehicle would overshoot 
the path of preceding vehicle and the MUCH system train 
would not track parallel to the section belt. The over- 
shooting problem was solved by increasing the steering 
bar length by 2 in. The problem of the MUCH train not 
tracking parallel to the section belt was corrected with 
addition of hydraulic power for steering the discharge vehi- 
cle. The hydraulic power pack was added to the bridge 
conveyor, and the control valve and steering cylinder were 
added to the discharge vehicle. Other problems, such as 
jackknifing, were not solved although many modifications 
and/or changes were made in the system. The results of 
1977-79 testing thus remained inconclusive. 

In 1980, JMMD was asked to conduct additional testing 
on the MUCH system. The objectives of additional testing 
were to find out the MUCH system limits, its ability to 
operate and stop on inclines, and how it would track 
through S-turns on various slopes, over the rolls, and over 
loose and mud bottom. These tests were conducted at the 
Ohio Transportation Research Center (OTRC). 

A three-entry mine plan was laid out at the OTRC. 
The simulated walls of entries and crosscuts were con- 
structed with snow fence. The system would not track- 
retrack in the same manner at the OTRC as it had at the 
Jeffrey facility. It required modifications and changes for 
steering and wheel alignment, as well as addition of turn- 
buckles to all vehicles and brakes on the lead and dis- 
charge vehicles. After these modifications and changes 



were made, there was an improvement in track-retracking 
and conveyor discharge performance. 

In 1982, additional design changes were made prior to 
sending the system to the Bureau's test facility at Bruce- 
ton, PA. These changes included installation of a chain 
between hopper and conveyor to limit travel between the 
adjacent vehicles, installation of rubber belting on hopper 
sideboards to stop spillage, and relocation of the conveyor 
speed switch. 

METF TEST PROGRAM OVERVIEW 

The MUCH system, consisting of 12 vehicles, bridge 
conveyor, and numerous spare parts, was received at the 
Bureau's METF in July 1983 for assembly and surface 
testing. Upon completion of assembly, all functions and 
safety devices were checked out and made operational; 
operator familiarization and training was undertaken prior 
to starting the test program. 

Surface tests were conducted at the METF to verify and 
evaluate the performance of the MUCH system. Tests 
were divided into sequences to evaluate a particular sub- 
system or machine function. Modifications were made to 
the MUCH system to correct deficiencies noted during 
surface testing. A description of each test sequence is 
given in the following sections. 

TRAM AND MANEUVERABILITY TRIALS 

Tests were conducted in the METF equipment maneu- 
verability trial area (EMTA) to determine the tramming 
capability of the system in a simulated mine environment, 
to define and correct any observed tramming or steering 
problems, and to demonstrate the reliability of the overall 
tram system. 



Early in the surface test program, it became obvious 
that tramming and maneuvering the 12-vehicle system was 
no simple matter, especially within the confines of the 
EMTA and with inexperienced operators. The system was 
trammed into a continuous loop numerous times for haul- 
age system trials, demonstrations, and noise level tests, 
and was also trammed in and through the simulated mine 
workings of the EMTA. 

Figure 6 illustrates a number of operational procedures 
and maneuvering sequences that were undertaken early in 
the tramming trials. Initially, these maneuvers were diffi- 
cult to perform successfully. Vehicles were contacting the 
ribs at corners as well as midrib. After operators per- 
formed a number of maneuvers, they were better able to 
gauge how an entry or crosscut must be entered, how tight 
a radius to turn, when to begin straightening the lead vehi- 
cle in a turn, how much rib clearance was needed, etc. 
Figures 6B through 6F illustrate typical maneuvers per- 
formed to place change the system from a far-right entry 
to the left crosscut in a simulated three-entry section. 
Figures 6F through 6H illustrate the return of the system 
to the far-right entry. In the step shown in figure 6£>, it 
was necessary to reposition the outby end of the system to 
provide sufficient room to accommodate the limited turn- 
ing ability of the discharge vehicle. 

During these early tramming trials, more problems 
were experienced with the system making rib contact near 
the center of the pillar (area 1, fig. 6H) than at the pillar 
corners (area 2, fig. 6H). At the pillar corners, when 
making a 90° turn, the system tended to move away from 
the corner, but at midpillar, the system worked toward the 
rib and sometimes made contact. In general, the overall 
tracking of the system seemed inconsistent, but the track- 
ing seemed better when turning sharp 90° turns than when 
making more gentle, larger radius turns. 

Figure 67 shows the system in a gentle S-curve being 
trammed through the EMTA and out the east equipment 
door. This maneuver proved to be one of the most diffi- 
cult encountered during the early tramming trials. Three 
pillar corners and the exit door frame were contacted 
numerous times while the system was being trammed out 
of the building. 

It was much more difficult to move the system away 
from the rib or corner when the train was relatively 
straight than when the train was in a tight turn. This was 
expected, because the force available to pull a vehicle 
away from a rib or corner is proportional to the sine of 
the angle between the jammed and adjacent vehicle, and 
dependent upon the radius of curvature of the train. 

The typical method used to unjam vehicles from contact 
with a pillar was only marginally effective when the train 
was in a gentle curve. The method consists of shutting off 
outby vehicles and tramming the remaining vehicles inby, 
then pulling vehicles away from the rib. Likewise, unjam- 
ming vehicles from contact with a pillar corner by remov- 
ing tram power from outby vehicles and tramming remain- 
ing vehicles in the outby direction to push jammed vehicles 
away from the rib was ineffective when the train was in a 



gentle curve. These methods did work quite well, how- 
ever, when the system was in a sharp turn. 

A coal ramp approximately 2.5 ft high (fig. 6/) was 
built to observe system performance as vehicles trammed 
over it. The lead vehicle and first two intermediate 
vehicles were successfully trammed over the coal ramp and 
down the other side despite deep wheel penetration. No 
serious mechanical interferences were observed. Steering 
bars provided the necessary degrees of freedom in roll, 
pitch, and yaw axes. As the overall test program 
continued, a vehicle-to-vehicle interference problem that 
led to major vehicle frame modifications became obvious. 
This problem and subsequent modifications are described 
in the "System Modifications" section of this report. 

As a result of these initial tramming trials, the steering- 
tracking system was examined to better in defining the 
theoretical capabilities of the system and correct any 
mechanical problems which might hinder tracking ability. 

Steering System 

A physical layout of the mechanical steering system 
(fig. 7) was constructed that reflected the linkage geometry 
of one steering axle and drawbar. This model led to a 
better understanding of the steering system and helped in 
defining the theoretical limits of the system. Figure 8 
shows an overall layout of the system making two turns of 
different radii. During tramming of the system, it had 
been observed that the tracking seemed to be better when 
making tight turns than when making more gradual turns. 
The tracking layout drawing supports this observation. It 
can be seen that when the system is making a 12-ft-radius 
turn, the center of the turn is coincident for all vehicles 
making the turn. 

A 43-ft turning radius is also shown in figure 8. It can 
be seen that the center of the turn wanders as the turn is 
entered by successive vehicles. It moves approximately 
14 in with each new vehicle entering the turn. An en- 
larged view (fig. 9) shows the tracking error of each vehi- 
cle as the vehicles pass the same reference point. A track- 
ing error of 5.25 in per vehicle occurs when turning at a 6° 
drawbar angle, which yields a 43-ft turning radius. 

Once the tracking was defined from a theoretical stand- 
point, it was obvious that the system was not tracking as 
well as could be expected; therefore, the mechanical steer- 
ing components of each vehicle were examined. The steer- 
ing system freeplay was measured with a dial indicator on 
the end of each steering bar where it attached to the tie 
rods. Freeplay ranged from 0.03 to 0.050 in. The causes 
of excessive freeplay were found to be two loose tie rods, 
which were tightened, and 26 loose steering plates on 14 
axles. The steering plates are bolted to the steering 
knuckles by four bolts on each plate. Once the bolts 
loosen, the plate is free to rotate as much as the clearance 
in the boltholes will allow, thus allowing excessive freeplay 
in the system. All loose steering plate bolts were torqued 
to specifications. 



*48'-»H20W 

20' 

"T 

48' 



n: 



\ 



r 



Lead vehicle 



Continuous loop 
haulage demonstration 



]oD 



D 



Tram into entry 



«3 



..J 



D 

D 



□ 



g- 



— o 
D 



Tram ahead to provide room 
to back up and steer discharge 
car 



• ( 



» 



D 

D 



Tram out by 



f 


V 


1 





n 



i 



l \ / Tram outby to place 

change 



□ 



Dl 



^ 



!<*&- 



H 



D 

D 



Begin simulated 
place change 



JJO 



D 



..J 



Complete place change 
into new entry 



J % 



D 

D 



1 / Tram over coal pile 



I I 

Figure 6. -MUCH system operational maneuvers in equipment maneuverability trial area. 



Drawbar 




Figure 7.-Mechanical steering linkage geometry. 



10 




Figure 8. -Tracking layout, 12- and 43-ft turning radii. 



The mechanical steering system on each vehicle was 
properly aligned per instructions in the operator's manual. 
Adjacent vehicles were trammed into a straight line and 
the drawbars were aligned along the axis of the vehicles. 
The tie rods were then adjusted so that each wheel was 
parallel to the drawbar center line. 

The elimination of excessive freeplay in the steering 
system and the alignment of the individual wheels had a 
positive effect on the system tracking, but some problems 
were still evident, especially when tramming in the outby 
direction. Tramming outby was difficult because of the 
severely limited steering range of the discharge vehicle 
and the tendency of the vehicles in the trailing half of the 
system to track poorly and jackknife. Modifications were 
made to the discharge vehicle steering system to increase 
the available steering angles and improve component 
location. The modifications are described in the "System 
Modifications" section of this report. 

The tendency of the trailing half of the system to track 
poorly and jackknife while tramming in the outby direction 
was investigated and the problem was corrected. The tram 
system is designed so that the electrical power to the tram 
motors in the last two trailing vehicles is automatically cut 



off during tramming to provide tension in the train. When 
the tram control in the lead vehicle is actuated, all tram 
motors receive power, but after approximately 1.1 s, the 
motors in the two trailing vehicles are automatically pow- 
ered down to create drag on the vehicle train. 

Upon investigation, a malfunctioning time-delay relay 
was found in the lead vehicle outby tram circuit that did 
not power the motor down after 1.1 s of operation. The 
unit was disassembled and repaired. Subsequently, track- 
ing in the outby direction improved and jackknifing when 
tramming outby was reduced, because additional drag was 
provided by the lead vehicle. 



Tracking-Retracking Tests 



Objective 



The objective of the tracking-retracking tests was to 
determine the ability of the MUCH system to successfully 
tram in both the inby and outby directions within the con- 
straints of the EMTA simulated workings. 



11 




Direction 
of travel 



Approx A _ B tracking 
misalignment =5/4 in 



Figure 9.-Tracking capability, 43-ft turning radius. 



Procedure 

The EMTA was utilized to simulate an underground 
working area with 48- by 48-ft pillars and 20-ft-wide entries 
and crosscuts, as shown in figure 10. Ten data stations 
were located on the test course. Each data station con- 
sisted of a heavy string pulled taut across the entry approx- 
imately 5 ft above the floor. A 20-ft-long by 1-ft-wide strip 
of heavy kraft paper was suspended from each string. 
Uncapped felt-tip markers of assorted colors were attached 
to each of the 12 MUCH system vehicles, one per vehicle, 
at the same location on each vehicle. As the system was 
trammed through the test course, the markers were drawn 



across the kraft paper at each data station to permanently 
record the relative position of each vehicle at each station. 
To conduct the tests, the entire MUCH system was 
trammed through the course of 10 data stations four times, 
twice in the inby direction and twice in the outby direc- 
tion. After each traverse, the positions of both the lead 
and discharge vehicle, as indicated by the associated felt- 
tip mark on the kraft paper at each data station, were 
measured relative to both the left and right ribs. While 
tramming in the inby direction, the system was controlled 
by the steering of the lead vehicle. In the outby tram 
direction, the discharge vehicle steering controlled the 
system. 



12 





KEY 

~ j Simulated ribs 



System tracking 
variance 

10 Data stations 



Figure 10.-Trial A (top) and trial B (bottom) inby. 



13 



Results and Discussion 

The results of the tracking-retracking tests are given in 
table 1 and are shown in figures 10 and 11. Table 1 pre- 
sents the analytical data acquired during the trials. Fig- 
ures 10 and 11 present layout drawings of the test course 
that show the dimensions of the course, data station loca- 
tions, direction of system travel, path of the operator- 
controlled vehicle through the course, and tracking error 
(variance) of the system through the course. 

In table 1, the operating width column shows the max- 
imum width required by the system at the associated data 
station. This width is equal to the width of the operator- 
controlled vehicle (6 ft 8 in) plus the maximum amount of 
tracking error (variance) shown by any other vehicle at the 
data station. The maximum variance column shows the 
maximum variance observed at each data station and 
whether the variance was to the left or the right of the 
position of the operator-controlled vehicle. The rib dis- 
tance column shows the positioning of the required operat- 
ing width of the system in reference to the left and right 
ribs at the data station. The left and right designations are 
relative to the operator position, which is always with the 
first vehicle facing the direction of travel. 

During the tramming trials, the MUCH system tended 
to drift toward the inside radius of the turns. This tight- 
ening in the curves is probably due to the tramming resis- 
tance imposed by the automatic braking of the last two 
cars on the trailing end of the system. This braking keeps 
the system in tension, which keeps individual vehicles from 
jackknifing, but also tends to pull the vehicles toward the 
inside of a curve. The maximum variance that occurred 
while tramming in the inby direction was 68-in, which 
occurred at data station 5 during trial A (table 1, fig. 10). 

Tramming in the outby direction was more difficult 
than tramming inby because of the more limited steering 
capacity of the discharge vehicle. During both trials, rib 
contact was made at the corner where data stations 5 and 
6 intersect (fig. 11) when tramming in the outby direction. 
This corner was also the point of maximum system vari- 
ance, 94- and 101-in, during trials A and B, respectively. 
During trial A, intermediate vehicle 4 contacted the cor- 
ner. In trial B, as shown in figure 11, the system was 
trammed closer to the left rib prior to starting the turn to 
the right, and the turn was initiated sooner to create a 
smoother flowing curve. During this trial, only the last 
vehicle intersected the rib, although the overall variance 
was greater than in trial A. 

Because of the lightweight fiberglass panels in the 
EMTA, the rib contact during tramming was a problem 
that prevented further tramming until the system was 
moved away from the rib. In a similar underground situa- 
tion where rib contact would cause no damage, the system 
could have continued tramming while being guided by the 
rib. While rib contact is not desirable, it would not be as 
much of a problem underground as in the EMTA. In the 
author's opinion, the system tracked and retracked with 
enough consistency to operate satisfactorily in an under- 
ground mine of similar dimensions. 



Simulated Production Cycle Trial 

Objective 

The objective of the simulated production cycle trial 
was to evaluate the tramming and tracking ability of the 
MUCH system during a simulated face production cycle. 

Procedure 

The EMTA was utilized to simulate an underground 
working area with 48- by 48-ft pillars, 20-ft-wide entries, 
and 20-ft-wide, 90° crosscuts. A National Mine Service 
shuttle car was utilized to simulate a continuous miner at 
a production face. The simulated production face was in 
the center entry of the three-entry simulated workings. 
The discharge vehicle of the MUCH system was in the 
left-hand entry; the 10 intermediate vehicles ranged from 
the discharge vehicle, through a 90° right crosscut to the 
center entry, through 90° left-hand turn to the lead vehicle, 
which was positioned directly behind the simulated miner 
at the face (fig. 12). 

A continuous miner cut cycle was simulated by advanc- 
ing the simulated miner and MUCH system 2 ft to simu- 
late a sump and shear cycle, retreating 4 ft to prepare to 
cut the cusp and clean up, then advancing 6 ft to cut the 
cusp, clean up, and sump and shear. This cycle was re- 
peated 10 times to simulate a 20-ft face advance. Upon 
the completion of this cut cycle, the simulated miner and 
MUCH system were backed up approximately 30 ft, the 
simulated miner was repositioned to the left-hand side of 
the face, and the cut cycle was repeated on the second 
lift. The overall functioning of the tram system, system 
alignment, and tire tracks were observed during the simu- 
lated cutting cycles. 

Upon the completion of the simulated cutting cycles, 
the MUCH system was trammed to simulate a place 
change. The MUCH was trammed outby, down the mid- 
dle entry, through the crosscut, and back into the left- 
hand entry. The system was then trammed straight ahead 
in the left-hand entry, past the crosscut and inby into the 
next 90° crosscut, and to the right to simulate a place 
change. The time required to complete with place change 
was measured. 

Results and Discussion 

The MUCH system trammed without problem and 
tracked fairly well during these trials. No rib contact, jack- 
knifing, or tramming delays occurred during tests. The 
tracking of the system during the simulated cut cycles was 
good. As was evidenced in the previous tracking-retrack- 
ing tests, the system tends to wander toward the inside 
radius of the 90° turns. This tendency was also observed 
during these trials but to a lesser degree. 

The maximum tracking variance observed during these 
trials is shown in figure 13. This series of tire tracks 
shows the tracking variance at the 90° turn located at 
the intersection of the crosscut and middle entry. This 



14 



"footprint" is the accumulation of tire tracks created 
while the system was trammed inby to the face, operated 
through a 20-ft cutting advance, a lift change, a second 20- 
ft cut advance, and then trammed outby during the place 
change. The total width of the track is 35 in. Therefore, 



the MUCH system required a total working width of 35 in 
plus the width of the lead vehicle to operate without rib 
contact. The total required operational width would be 
9 ft 7 in at that location. 








KEY 
I | Simulated ribs 

MM System tracking 
variance 

I -10 Data stations 



Figure 1 1 -Trial A (top) and trial B (bottom) outby. 



15 



Table 1.-MUCH tracking-retracking trials 



Data 


Operating 
width 


Max variance, in 






Rib distance 






station 


Left Right 




Left 




Riqht 






ft in 


ft 




in ft 




in 



Trial A, in by: 

1 

2 

3 

4 

5 

6 

7 

8 

9 

10 

Trial B, inby: 

1 

2 

3 

4 

5 

6 

7 

8 

9 

10 

Trial A, outby: 

10 

9 

8 

7 

6 

5 

4 

3 

2 

1 

Trial B, outby: 

10 

9 

8 

7 

6 

5 

4 

3 

2 

1 

*Made contact with rib. 



8 

7 
10 
10 
12 
10 

8 

8 
10 

9 

9 
9 

7 
8 
9 
9 

7 
8 
7 
9 

9 
8 
9 
8 
12 
13 
8 
8 
8 
8 

9 

8 

8 

10 

14 

11 

9 

8 

8 

11 



8 

10 
6 
7 
4 

2 
4 
4 
8 

5 

7 

7 

1 

10 

10 

2 

6 

11 



7 
10 
4 
9 
1 
10 
8 
6 
9 
8 

7 
2 
8 
8 
4 
3 
3 
10 
11 
2 



24 

14 

46 

47 

68 

40 

8 







33 

35 

11 

17 

36 

36 









44 

34 

26 











33 

32 

44 

26 

32 











35 

61 















10 

20 

44 

36 













6 

22 

15 

28 





14 

33 

86 

94 

32 

30 











55 

101 

70 

39 

34 







6 
5 
1 
1 
2 
3 
6 
8 
7 
6 

4 
4 
6 
6 

4 

7 
6 
7 
6 
6 

1 
1 
4 
10 
7 
6 
9 
6 
4 


1 
2 
4 
6 
5 
8 
10 
9 
5 
1 



4 

10 
9 
9 
3 

11 
8 

10 


11 

8 
5 

4 
9 
9 
3 
5 
4 
5 

1 
7 

10 
2 

11 
2 
7 

11 
3 

11 

5 

6 

10 

7 
8 
9 
2 
7 
5 
3 



5 

6 
7 
7 
5 
6 
5 
2 
2 
3 

5 
6 
6 
5 
5 
2 
6 
4 
5 
4 

9 
9 
5 

1 



1 

4 

7 

10 

9 
9 
6 
2 





1 

5 
6 




4 
9 
8 
5 
1 
2 
10 
8 
5 

11 

5 

7 
5 
5 

7 
1 
9 

7 

4 

7 

10 

1 

J 
'0 
9 
7 

5 


4 
6 

:? 

*o 

7 

7 

8 

11 




Figure 12.-Lead vehicle behind shuttle car. 



16 




Conveyor Noise-Level Survey 



Objective 



Figure 13.-Tracking variance. 

The simulated place change that involved tramming the 
system outby from the middle entry, through a crosscut 
into the left-hand entry, then inby in the left-hand entry to 
the next crosscut and turning into the crosscut, covered a 
total tram distance of 245 ft. The outby tramming for the 
place change covered a distance of 155 ft and required a 
total time of 1 min 30 s, there was a pause of 10 s while 
the lead vehicle operator communicated with the discharge 
vehicle operator before starting to tram inby to complete 
the place change. The inby tramming covered a distance 
of 130 ft and required 1 min 40 s. The complete place 
change required a total of 3 min 20 s to complete and no 
problems occurred during the trial. 

CONVEYOR SYSTEM TESTS 

The main purpose of the MUCH system is to continu- 
ously convey coal cut by a continuous miner at the face to 
the section panel belt. Therefore, a large portion of the 
MUCH test program was devoted to testing, evaluation, 
repairs, and modification of the conveyor system. 



A conveyor noise level survey was conducted to deter- 
mine noise exposure experienced by operators or miners 
working close to the MUCH system. 

Procedure 

The system was set up in a continuous loop behind the 
EMTA. The train was configured in a circle with the 
discharge vehicle dumping into the lead vehicle hopper 
while operating in the conveyor mode only. Tests were 
performed with the conveyors empty and with varying 
amounts of coal in the conveyors. Coal was a mixture of 
2.5- by 2-in and 2- by 1.5-in sizes. All tests, except ambi- 
ent noise measurement, were performed with the convey- 
ing system in operation. 

Noise-level measurements were obtained in the center 
of the vehicle circle, around the outer perimeter of the 
vehicle train, and in the operator compartment. Attention 
was focused on the operator compartment of the lead 
vehicle. 

A Bruel and Kjaer type 2205 handheld sound level 
meter (SLM) with a Bruel and Kjaer type 4117 piezo- 
electric microphone was used to perform the sound level 
measurements. Sound pressure level (SPL) was measured 
using the A-weighted network. This filter network patterns 
its response after the human ear. SPL meter response was 
in the slow mode. 

All measurements were taken with the meter held away 
from the body to minimize the effects of noise reflection 
from the body. It should be noted, however, that perform- 
ing a noise survey in an enclosed building, particularly with 
a large number of hard surfaces reverberating or reflecting 
noise, may cause SPL measurement errors. Larger objects 
with physical dimensions similar to the wavelength of the 
sound being measured are most likely to reflect noise and 
to be sources of error. Considering these limitations, the 
test area was adequate to perform such a general noise 
survey. 

Results and Discussion 

The noise level test configuration is shown in figure 14 
and the survey results are given in table 2. The three tests 
are discussed in the following sections. 

Test 1-Conveyor Off, Ambient Noise Level Measurement 

The measured ambient noise level of 50 to 51 dBA in 
the test area introduced no error into the operating noise 
level measured. Ambient noise level greater than 10 dBA 
below the measured SPL will introduce no significant error 
in the data; therefore, no compensation for ambient noise 
level was made. 



17 



Table 2.-Conveyor noise test results (sound pressure levels) 

Configuration 

Conveyor off, ambient noise level 

Conveyor operating, no coal in conveyor: 

Reading from center of conveyor loop (sound level meter rotated 360°) 

Pass around (except at lead and discharge vehicle junction) 

Junction of lead and discharge vehicle 

Operator's ear level in lead vehicle 

Conveyor operating with coal, level at lead vehicle operator's ear: 

Approximately 3.7 st/min 

Approximately 7.4 st/min 

Approximately 11.1 st/min 



dBA 



50 - 51 





98.3 


101 


-102 


103 


-104 




102 




98 




97 


96.5 


- 97.5 



KEY 
Sound level meter 

EMTA Equipment maneuverability trial area 



EMTA 



Discharge vehicle 




Lead vehicle 
Operator compartment 



Figure 14.-Conveyor noise-level test configuration. 



18 



Test 2— Conveyor System On, No Coal 

a. The SPL was measured in a 360° arc from a position 
in the center of the conveyor loop (fig. 14). The SLM was 
held away from the technician's body to minimize error, 4 
to 5 ft above the ground and pointing toward the outside 
of the vehicle train circle. The measured SPL was 98.3 
dBA. 

b. A pass-by noise measurement was made by walking 
parallel to the train on the outside perimeter, approxi- 
mately 2 ft from the vehicles. The SLM was held 4 to 5 ft 
above the floor and away from technician's body. The axis 
of the SLM was held parallel to the tangent of the MUCH 
circle, approximately 2 ft away from the vehicles. The 
SLM was held parallel to the vehicles to maintain geo- 
metric uniformity and to facilitate meter reading. The 
pass-by noise measurement is most closely associated with 
the noise level that a nearby miner would experience when 
in the same entry with the system. The measured SPL 
ranged from 101 to 102 dBA. 

c. The SPL measured at the lead-discharge vehicle 
junction was 103 to 104 dBA in the pass-by mode as in test 
2b. It was slightly more noisy than the junction of any 
other two vehicles. 

d. The SPL in the operator's compartment at the 
operator's ear level was measured to determine noise level 
exposure of the lead vehicle operator. The operator com- 
partment, at 102 dBA, was very noisy. Moving the micro- 
phone around the vicinity of the operator's ear showed 
minimal change in the readings, and thus increased the 
confidence in the SPL measurement. Accurate meter 
positioning was, therefore, found unnecessary. 

In addition to customary hearing protection, a head- 
set for communicating with the discharge vehicle opera- 
tor would significantly attenuate the noise level to the 
operator. 

Test 3-Conveyor System On, With Coal 

A front-end loader was used to meter approximately 
3 yd 3 of coal onto the MUCH system. The SPL at opera- 
tor's ear level was measured while the system was con- 
veying coal at approximately 3.7 st/min. The introduction 
of coal into the conveying system, even in small amounts, 
significantly reduced the noise level by damping the con- 
veyor structure and absorbing acoustic energy. Noise level 
at the operator's ear was reduced by 4 dBA to 98 dBA. 

The test was repeated, this time with 6 yd 3 of coal in 
the conveying system (7.4-st/min haulage rate). A 1-dBA 
noise level reduction was achieved by this doubling of the 
amount of coal on the conveyor (from 98 to 97 dBA). 

The test was again repeated with 9 yd 3 of coal in the 
conveyors, an 11.1-st/min haulage rate. No further reduc- 
tion in noise level over the 6-yd test was achieved by the 
addition of 50 pet more coal (SPL, 96.5-97.5 dBA). It 



appeared that adding coal to the conveying system signifi- 
cantly reduced operator noise exposure, but further reduc- 
tion in noise level was minimal beyond the 3.7-st/min 
haulage rate. 

The noise levels measured indicate that some form of 
hearing protection will be required for the MUCH system 
operator to maintain an 8-h noise exposure of 90 dBA or 
below. 



Conveyor Speed Test 



Objective 



A conveyor speed test was conducted to determine con- 
veyor chain speed and motor rotational speed under both 
loaded and unloaded conditions and to determine the 
transport speed of discrete particles at various loading 
rates. 

Procedure 

A Micronta 63-5009 digital stopwatch was used to mea- 
sure time required for the conveyor chain in each vehicle 
to make a complete cycle. When the system was loaded 
with coal, the conveyor chains were not visible; therefore, 
a General Radio 1531AB stroboscope was used to measure 
tailshaft rotational speed. After establishing a relationship 
between chain speed and tailshaft rotational speed, the 
chain speed was measured with a stroboscope. A stop- 
watch was also used to measure the transport speed of 
individual chunks of coal. 

Results and Discussion 

The conveyor surface speed was measured when empty 
and when loading 2 and 7 st/min (table 3). At st/min 
(empty), the time required for the conveyor chain to make 
one complete pass through the vehicle was measured with 
a stopwatch. Three trials were conducted for each vehicle 
in the empty condition. 

The conveyor chain lengths were 45 ft 5 in for the lead 
vehicle and 42 ft 5 in for all other vehicles. Dividing chain 
length by time yielded the average conveyor speed for 
each vehicle, which ranged from 270.7 to 273.7 ft/min, an 
average of 272.2 ft/min. The 272.2-ft/min average agrees 
within 2.7 pet of the published figure of 280 ft/min in the 
work cited in footnote 4. 

Conveyor tailshaft speed was also measured to deter- 
mine the speed relationship between the conveyor and 
tailshaft. Tailshaft rotational speed averaged 266.0 r/min 
under empty conditions. Thereafter, when coal was in the 
system, the conveyor speed was measured via tailshaft 
rotational speed. The ratio of conveyor speed to tailshaft 
rotational speed (K r/min ) was 1.025. Multiplying tailshaft 
rotational speed by K r/min gives the equivalent chain speed. 

At a conveying rate of 2 st/min, conveyor chain speed 
averaged 268.3 ft/min. At 7 st/min, average chain speed 
was 270.9 ft/min. From to 7 st/min, average conveyor 
speeds were within 1.6 pet of one another. At 2 st/min. 



19 



average conveyor speed was 2.6 ft/min slower than at 7 
st/min. This could be attributed to the small sample 
taken at 2 st/min (four vehicles) and the possibility of 
fines buildup on the conveyor deck during the trial. 

Transport speed of individual particles was measured at 
various loading rates. Several chunks of coal and coalcrete 
(from 2 to 6 in) were timed as they passed through the 
system (table 4). With no coal in the system, the average 
transport speed was 253.5 ft/min. This figure is about 7 
pet less than the average conveyor speed of 272.7 ft/min, 
reflecting the time required for the conveyor chain to pick 
up coal as it cascaded onto the next outby vehicle. 

With the system conveying about 2 st/min, transport 
speed was about 3.2 ft/min less, or 250.3 ft/min. At 4 
st/min, transport speed was 234.0 ft/min, but only one 
trial was performed. 

Generally, transport speed will decrease somewhat at 
higher loading rates. From observation, it can be seen 



that some coal tends to ride over the chain flights as load- 
ing rate increases, thus tending to decrease the transport 
rate. It was also observed that as coal became finer from 
repeated handling, a greater percentage of it tended to 
ride up over chain flights, further reducing average trans- 
port speed. With relatively large pieces (2 to 6 in), how- 
ever, no correlation could be seen between particle size 
and transport speed. 

Another way of interpreting transport speed is to relate 
it to conveyor speed, providing a measure of conveying 
efficiency. At st/min, transport speed is 253.5 ft/min 
and conveyor speed is 272.7 ft/min. Transport speed (or 
conveying efficiency) is 93 pet of the conveyor speed. Con- 
versely, slippage would only be 7 pet. A 100-pct efficient 
system would transport coal at the same speed as the con- 
veyor travels. At 2 st/min the effective conveyor transport 
efficiency is 93.3 pet, and at 4 st/min it is 87.3 pet. 



Table 3.-Conveyor speeds 





Time 


per convevor evele, 


s 


Tailshaft Av 
speed, conveyor speed, 1 


Conveyor speed 


Vehicle 


Trial 


Trial 


Trial 


to tailshaft ratio 




1 


2 


3 


r/min 


ft/min 


( K r/min)- r / min 






HAULAGE RATE: 


st/min 


(SYSTEM EMPTY)-3/28/84 






Lead 


10.04 


9.88 


10.02 


266 


273.1 


1.027 


1 


9.49 


9.31 


9.39 


265 


270.7 


1.022 


2 


9.34 


9.54 


9.30 


266 


271.0 


1.019 


3 


9.30 


9.33 


9.29 


266 


273.4 


1.028 


4 


9.27 


9.34 


9.35 


266 


273.1 


1.027 


5 


9.39 


9.31 


9.26 


266 


273.1 


1.027 


6 


9.27 


9.30 


9.32 


267 


273.7 


1.025 


7 


9.43 


9.37 


9.28 


266 


271.9 


1.022 


8 


9.31 


9.28 


9.30 


266 


273.7 


1.029 


9 


9.30 


9.34 


9.25 


265 


273.7 


1.033 


10 


9.31 


9.35 


9.35 


266 


272.5 


1.024 




9.26 


9.32 


9.42 


267 


272.8 


1.022 


Average 


NAp 


NAp 


NAp 


266.0 


272.7 


1.025 






HAULAGE RATE: 


2 st/min-3/14/84 






8 


NAp 


NAp 


NAp 


264 


270.6 


NAp 


9 


NAp 


NAp 


NAp 


257 


263.4 


NAp 


10 


NAp 


NAp 


NAp 


263 


269.6 


NAp 


Discharge 


NAp 


NAp 


NAp 


263 


269.6 


NAp 


Average 


NAp 


NAp 


NAp 


261.8 


268.3 


NAp 






HAULAGE RATE: 


7 st/min-3/30/84 








NAp 


NAp 


NAp 


262 


268.6 


NAp 


1 


NAp 


NAp 


NAp 


264 


270.6 


NAp 


2 


NAp 


NAp 


NAp 


264 


270.6 


NAp 


3 


NAp 


NAp 


NAp 


264 


270.6 


NAp 


4 


NAp 


NAp 


NAp 


264 


270.6 


NAp 


5 


NAp 


NAp 


NAp 


263 


269.6 


NAp 


6 


NAp 


NAp 


NAp 


263 


269.6 


NAp 


7 


NAp 


NAp 


NAp 


268 


274.7 


NAp 


8 


NAp 


NAp 


NAp 


266 


272.7 


NAp 


9 


NAp 


NAp 


NAp 


263 


269.6 


NAp 


10 


NAp 


NAp 


NAp 


266 


272.2 


NAp 


Discharge 


NAp 


NAp 


NAp 


264 


270.6 


NAp 


Av 


NAp 


NAp 


NAp 


264.3 


270.9 


NAp 



NAp Not applicable. 
'Calculated from total 



length of conveyor chain: lead vehicle-45 ft 5 in, all other vehicles-42 ft 5 in. 



20 



Table 4.-Particle transport speed 

Time. 1 s 

Particle diam, in Trial Trial Trial 

1 2 3 

0-st/min haulage: 

4 53.40 53.44 54.09 

4 54.69 NT NT 

6 53.87 54.55 54.89 

2-st/min haulage: 

2 57.10 NT NT 

5 53.02 54.99 NT 

6 54.35 NT NT 

4-st/min haulage: 

4 58.66 NT NJ_ 

NT No trial. 

transport time from lead vehicle hopper to end of discharge vehicle. 



Av 



Transport speed, ft/min 



Av for all sizes 
at loading rate 



255 

250. 
252. 



i} 



240.4 
254.1 
252.5 

234.0 



253.5 



250.3 



234.0 



Conveyor Time Sequencing 

Objective 

A conveyor time sequencing test was conducted to 
measure the startup sequence time required to start the 
conveyor system. 

Procedure 

Time required to power up all conveyors sequentially 
was recorded under loaded and unloaded conditions. 
During testing, with haulage rate established and coal 
evenly distributed throughout the MUCH system, the time 
required to restart the system was measured by the lead 
vehicle operator using a digital stopwatch. Time was 
measured from the time the conveyor start switch was 
pushed until the lead vehicle conveyor system started. 

Results and Discussion 

The time required to start up the entire train of convey- 
ors was recorded at haulage rates of 0, 2, 3, and 4 st/min. 
From data in table 5, it can be seen that the average 
startup ranged from 1.36 s at no load (0 st/min) to 1.86 s 
at 4 st/min, generally increasing as the load on the convey- 
or increased. 

On March 14, 1984, it took 2.73 s to start up the system 
at 2 st/min, an apparent anomaly. This was the first date 
of testing, and solidified coal had been in the idle convey- 
ors for a period of several days before the testing started. 
The additional load imposed by this material caused a 
significantly longer startup time. 

Coal Conveying Tests 

Numerous coal conveying tests were conducted during 
the program and all but the final test were followed by 
system modifications and improvements in an effort to 
establish the coal-conveying capability of the system and to 
achieve a 95-pct system availability while conveying at a 
rate of 8 st/min over an 8-h shift. 



March 14 to April 12, 1984 
Test Configuration 

A continuous haulage loop of the MUCH system was 
formed by having the discharge vehicle convey coal onto 
the 30-ft Long-Airdox belt structure and into the hopper- 
feeder-bolter (HFB). (Because of its ability to provide 
surge capacity, the HFB served as the coal-entry point for 
new coal being added to the system.) The HFB in turn 
dumped onto a 50-ft Long-Airdox belt structure, modified 
to accept the Ramsey Engineering belt weight scale, that 
loaded coal onto the lead vehicle of the MUCH system, 
thus completing the loop, as shown in figures 15 and 16. 

Coal used during the first part of testing was a mixture 
of 2.5- by 2-in and 2- by 1.5-in coal acquired from the 
Bureau's Hydraulic Transport Research Facility (HTRF). 
The second portion of conveyor testing was performed 
with run-of-mine (ROM) coal from the Bureau's research 
mine. 

Table 5. -Conveyor startup time 

Trial Date Startup time, s 

Sequenced Av 

0-st/min haulage: 

s ::::::::::::: [ ^^ \ !^ [ 136 

2-st/min haulage: 

1 "1 J2.83 "1 

I ::::::::::::: \ w* \ '?:§ \ 273 

4 J I 1.59 J 

!:::::::::::::} 3 ^ { IS 1 ,„ 

!:::::::::::::> ***« { !S J 

3-st/min haulage: 

1 I f 184 1 

2 > 4/03/84 < 1.81 > 1.87 

3 J I 195 J 

4-st/min haulage: 

2 ::::::;::::::} v™'** { I:* } 186 

'Measured during initial startup of conveyor testing. Solidified 
coal was present in conveyors, which increased startup time by a 
significant margin. 



21 




Figure 1 5.-Conveyor test configuration. 




MUCH, 
system 



Lead 
vehicle 



v. 





Conveyor 
drive gearbox 

Thermocouples 



Explosion- proof box^^ 
Transducer package 





HFB 



Belt 



Belt scale 



Belt scale 
electronics 



Thermocouple 
readout 



Recorder 



nstrumentation 
trailer 



Figure 1 6.-Test configuration and instrumentation system. 



22 



Instrumentation 

To more effectively evaluate the conveying system, 
pertinent electrical and mechanical parameters were mea- 
sured and recorded during conveying tests. Shown in 
figure 16 are the layout and location of the instrument 
systems. The following sections describe parameters mea- 
sured and instrumentation utilized. 

Total Conveying System Power 

Total system power was measured to provide an eval- 
uation of power requirements of the system, particularly 
under various operating conditions. 

True power, in kilowatts, of all 12 vehicles was mea- 
sured with a Rochester Instrument Systems watt trans- 
ducer. Current to the watt transducer was supplied by two 
400:5 current transducers, one on each of two phases of 
the incoming three-phase power. Current transformers 
were placed in the discharge vehicle's explosion-proof box 
on the load side of the main breaker, thus coupling trans- 
ducers to line power and reducing current by a factor of 
80. Voltage leads from the watt transducers were con- 
nected to the load side of the main breaker. 

Discharge Vehicle Power and Current 

Individual vehicle energy requirements are not neces- 
sarily one-twelfth of the total system power because of 
differences in instantaneous loading rate from one vehicle 
to another and also for the conditions that exists on a 
specific vehicle such as plugging or coal fines buildup on 
the conveyor deck. Therefore, the electrical power and 
current requirements of the discharge vehicle were singu- 
larly measured. A Rochester Instrument Systems watt 
transducer, identical to that measuring total system power, 
was used to measure the discharge vehicle power and a 
Transdata model 10CS501 current transducer was used to 
measure the current required by the discharge vehicle 
conveyor motor. Two 50:5 current transformers were 
placed in two of three conductors supplying power to the 
conveyor motor contactor, thereby coupling the trans- 
ducers to live current and reducing current by a factor of 
10. Voltage leads were connected to the load side of the 
main breaker. 

System Voltage 

To monitor regulation of the power center and to en- 
sure that the system operating voltage remained within 
acceptable limits, the system voltage was measured using 
a potential transformer and a potential transducer. The 
550:120 voltage tap was used on the Trenco TR12182 
potential transformer to supply an acceptable voltage level 
to a Rochester Instrument Systems 10PS101. The trans- 
former primary was connected directly across the line on 
the load side of the main breaker. Output of this potential 
transducer was proportional to system voltage. 



Haulage Rate 

Perhaps the most important variable measured in the 
test program was haulage rate. It is the independent vari- 
able to which most other dependent variables are related. 

A Ramsey Engineering, model No. 10-20/40-20, belt 
weigh scale was used to measure instantaneous loading 
rate. Belt-scale electronics included an integrator to total 
the number of tons of coal that was conveyed. 

Conveyor-Reducer Drive Temperature 

Lubricant temperature of two conveyor-reducer drive 
gearboxes was measured to evaluate suitability of the gear- 
box for loads imposed on the system. Two thermocouple 
junctions were made from Type K thermocouple wire. 
Each was entered through a sealed fitting into the gearbox 
sump in the discharge and the 10th intermediate vehicles. 
An Omega model 2168 digital thermocouple readout was 
used to measure gearbox temperature. 

A Gould model 481 eight-channel strip-chart recorder 
was used to simultaneously record total system power, 
discharge vehicle power and current, system voltage, and 
haulage rate. 

Procedure 

Instrument Calibrations 

All sensors being recorded on the Gould strip-chart 
recorder were physically calibrated prior to the test 
program. In addition, prior to each day's run, the Gould 
recorder was calibrated with approximate voltages for each 
channel to simulate proper sensor stimuli using a General 
Resistance DAS66AX Dial-a-Source. 

Physical calibrations of the current, potential, and watt 
transducers were performed using a calibration device con- 
sisting of a bank of resistors (unity power factor) arranged 
to provide three-phase voltage and current to the trans- 
ducers. Voltage devices were connected directly across 
the line. Current was measured using a Fluke 80J-10 
current shunt and Fluke 8600 DMM (digital multimeter). 
Voltage was also measured with the Fluke 8600 DMM. 
Transducer output was adjusted to the proper voltage level 
appropriate for input stimuli. Gain of the Gould recorder 
was adjusted to give appropriate scale deflection. These 
transducer voltages were inserted into the Gould recorder, 
via the Dial-a-Source, on a daily basis during the test 
program to maintain recorder calibration. 

The Ramsey belt scale was calibrated using weigh 
chains designed for that purpose. Calibration was per- 
formed with the belt running at normal operating speed 
with weigh chains placed on the moving belt. A three- 
point calibration was performed at 0-, 20-, and 40-pct full 
scale (0-, 4-, and 8-st/min loading rate). Gould recorder 
gain was also set to the appropriate level to indicate 
proper haulage rate. As with other transducers, the 



23 



recorder was calibrated daily using a simulated signal from 
the Dial-a-Source. 

Thermocouples were checked at ice point (32° F) using 
an ice bath and at boiling point (212° F) using a hotplate 
and water. 

Haulage Test 

A mixture of 2.5- by 2-in and 2- by 1.5-in coal from 
the HTRF was used in conveying tests from March 14 to 
April 5, 1984. For tests conducted April 5-12, 1984, ROM 
coal from the Bureau research mine was utilized. 

On March 14, 1984, first trial of the haulage test was 
performed. A Clark front-end loader was used to load 
coal into the HFB. Steady-state loading rate was adjusted 
from to 6 st/min over the test period from March 14 
through April 12, 1984. 

Total system power, discharge vehicle conveyor motor 
current and power, system voltage, and loading rate were 
recorded on the Gould strip-chart recorder for all loading 
trials. 

On each day the testing was performed, the Gould 
recorder was allowed to warm up for 30 min and then it 
was calibrated with the Dial-a-Source. Belt-scale totalizer 
readings were recorded at the beginning and at the end of 
testing to determine the total amount of coal transported 
each day. A log was kept of all significant events, break- 
downs, and repairs during the test period. The log was 
voice-recorded on a microcassette tape recorder and later 
transcribed. 

Conveyor gearbox temperature was measured before 
the start of each test and every 10 min thereafter for the 
first 30 min. Subsequent readings were taken every 30 min 
or whenever the system was shut down. 

After completion of testing, coal carryback loss lying 
beneath each vehicle's conveyor drive shaft was weighed 
with a 1,000-lb capacity balance beam platform scale. 

Results and Discussion 

One objective of the MUCH system test plan was to 
demonstrate reliability of the conveying system by operat- 
ing for one full shift at 8 st/min, with no more than 24 min 
of downtime (95 pet availability). Because of a variety of 
recurring problems, no more than 136 min of running time 
was achieved in any one day (28 pet availability). Table 6 
summarizes the amount of test time at the listed haulage 
rates. 



Table 6.-Haulage 

Av haulage rate, 
st/min 

(system emptv^ 


test 


summary 


Tesf time, 
min 

197.0 


1-2 








244.2 


3 








65.0 


4 








23.9 


5 








29.0 


6 








1.5 


Total 








'seo.e 


! 9 h 20.6 min. 











The maximum average conveying rate reached was 6 
st/min. Conveyor tests were performed in 13 days from 
March 14 to April 12, 1984. Total test time was 560.6 
min. Total run time achieved above 4 st/min was 30.5 
min. During the test period, 682.1 st of coal was con- 
veyed. Average loading rate for the entire period was 1.22 
st/min. Appendix B contains a conveyor test breakdown 
and repair log for this test period. 

The system test was started on March 14, 1984, at 2 
st/min, but after 2 days, only 65 min of run time had been 
achieved because of breakdowns. The system was then 
allowed to run empty for 2 days to determine if failures 
were load dependent. After the system had cleaned itself 
out, the breakdowns ceased, but then it took a period of 
time for the coal to be cleaned out from beneath conveyor 
decks. It appeared that after rehandling the coal many 
times and by adding water with the water sprays, the coal 
became a fine, granular, dense mixture, which required 
considerably more energy to convey than dry, coarse coal. 
This wet mixture plugged up around conveyor sprockets, 
under the conveyor deck, and even on top of the conveyor 
deck. 

As coal was again added to the system, mechanical and 
electrical problems began to reoccur. Even as the loading 
rate was kept constant at 2 st/min, the rate of system 
malfunction increased with time. It is hypothesized that 
this was due to increased buildup of wet fines everywhere 
in the system. 

Shown in table 7 is a summary of results for these con- 
veyor tests. Discharge vehicle current, discharge vehicle 
power, and total system power are shown for loading rates 
from to 5 st/min. Figure 17 shows typical Gould strip- 
chart traces of real-time data. 

System power consumption ranged from 45 kW (60.3 
hp) at no load (0 st/min) to 166 kW (222.5 hp) at 5 
st/min. Average power consumption per vehicle 3.8 kW 
(5 hp) and 13.8 kW (18.4 hp) at and 5 st/min, respec- 
tively. Measurement of discharge vehicle power indicated 
that its power consumption ranged from 7.5 kW (10.1 hp) 
at no load to 10.0 kW (13.4 hp) at 2.2 st/min. This was 
greater than the discharge vehicle power measured at 5 
st/min (8.9 kW or 11.9 hp), because wetted coal fines had 
built up on the conveyor deck to a depth of 4.5 in. Fig- 
ure 18 shows the extent of fines buildup on the conveyor 
deck. This buildup increased conveyor chain tension sig- 
nificantly and increased the coefficient of friction, requir- 
ing much greater power to convey the coal. Such buildup 
on the conveyor deck was also a problem with intermedi- 
ate vehicle 5. 

Once the coal fines buildup was removed from the 
deck, discharge vehicle power went down to 7.8 kW (10.5 
hp), a reduction of 22 pet. Likewise, discharge motor 
current decreased from 21 to 14.8 A (25 pet decrease). 

Under no-load conditions, the system consumed up to 
36 pet of the power required at 5 st/min. This was due 
mainly to the wet fines plugging up the conveying system, 
particularly material carried back under the conveyors in 
the vehicle hoppers. On March 16, 1984, after the system 
became plugged with coal and was emptied, residual 



24 



material caused the power consumption to remain 
artificially high at 68 kW (91.1 hp), even though no coal 
was being conveyed. On March 20, 1984, after allowing 
the system to run a period of time to purge itself, the 
power dropped 18 pet to 56 kW (75 hp). At that point, 
water sprays were turned on. After wetting the conveyor 
chains and deck for approximately 5 min, power 
consumption dropped another 18 pet to 46 kW (61.7 hp). 
One measure of conveying efficiency is the amount of 
energy required to convey material a given distance, 
expressed as specific energy in kilowatt hours per short 
ton. 



E *p = 



where E^ = specific energy, kW • h/st, 
P, = total system power, kW, 



and 



O = loading rate, st/min. 



O(60) 



In terms of energy requirements, most efficient 
conveying is achieved at the higher loading rates. At 4 
st/min, specific energy was 0.41 kW • h/st conveyed. At 
the 1.8-st/min loading rate on March 15, 1984, the specific 
energy was 1.36 kW • h/st. Poorer efficiency was due to 
two factors: (1) material had built up on conveyor decks 
and underneath conveyors, and (2) there is a certain tare 
power requirement to overcome system friction under no- 
load conditions. 



Table 7. -MUCH haulage testing data summary-electrical parameters 



Haulage rate, 
st/min 


Discharqe convevor 
Current, Power, 
A kW 


System 

power, 

kW 


Total 

system 

voltage, 

V 


Coal conveying 

specific energy, 1 

kW-h/(st-ft) 


Comments 


3/14/84: 
2.4 . . 




13.5 

15.5 
13.0 
17.7 

17.0 
14.0 
19.8 

15.2 


NA 

NA 
NA 
NA 

NA 
NA 
NA 

8.4 


108 
117 

147 
118 


465 
465 

460 
465 


0.75 
.85 

1.36 
1.23 


Material had been on conveyor 
decks for several weeks prior 
to system startup. 


2.3 . . 




1.0 . . 




Values averaged over data 


3.1 . . 




window. 


3/15/84: 
1.8 . . 




Material buildup on conveyor 
decks, mechanical problems 
occurred because of high loads. 

15.5 st coal conveyed. 


0.7 . . 




4.0 . . 




3/16/84: 


1.6 


3/20/84: 
. . . 
.. . 




11.8 
12.2 


7.8 

8.0 


68 
60 


475 
480 


NAp 
NAp 


At 2: 11 p.m. 
At 3:11 p.m. 


3/28/84: 
.. . 
... 
. . . 




12.0 
10.2 
12.2 


7.8 
7.5 
7.8 


56 
46 
45 


485 
485 
485 


NAp 
NAp 
NAp 


Before addition of water, 2:51 p m 
After addition of water, 2:56 p.m. 
At 3:14 p.m. 


3/29/84: 


2.0 


13.0 


7.6 


87 


475 


.73 


38.3 st coal conveyed. 


3/30/84: 
2.0 . . 




11.5 

13.0 

21.0 


7.4 
7.4 

10.0 


81 
99 

138 


475 
470 

460 


.68 
.41 

1.05 




4.0 . . 




196.8 st coal conveyed. 


4/05/84: 


2.2 


Discharge and intermediate vehi- 
cle 5 conveyor decks were 
plugged. 


4/06/84: 


3.0 


14.8 


10.5 


138 


460 


.77 


After-discharge and intermediate 
vehicle 5 conveyor decks cleaned. 
59.8 st coal conveyed. 


4/09/84: 


5.0 


18.0 


8.9 


166 


455 


.55 


81.7 st coal conveyed. 


4/12/84: 





NAp 


NAp 


NAp 


NAp 


NAp 


Mechanical failures and electrical 
problems effected the recorded 
data. 24.9 st coal conveyed. 



NA Not available. 

NAp Not applicable. 

^otal mean-0.84 kWh/(st-ft); standard deviation, 0.30 kWh/(stft). 



NOTE.-ln tests conducted April 3 (44.9 st coal conveyed) and April 4 (46.6 st coal conveyed), electrical and mechanical problems were 
encountered. No operating data were obtained. 



25 



-to. 

LdCE 

gee 

ozO 
</>0°- 

c?£z 

ok 

sco?, 

QO 



o 



£&> 

UJ 

x 



150 

100 

50 


20 

10 



250 
12.5, 



March 14, 1984 



Chart speed = 0.5 mm/s 



■»%»i ■ to. 



.000 



500 - 



20 
J 10 



— I ■ 1 1 1- 

Ji'u-* • _ *~ 1 * ' 



_L 



i 1 1 1 . 

Discharge conveyor power no 



100 



operating 




200 



300 



400 500 

TIME, s 



600 



700 



800 



900 



war . 

UZO 
QU 

ujrr.- 

£££< 
SOU 



| -R> 

to -I 

>-o 
to> 



UJ 



^3. 



300 
200 
100 







March 30, 1984 



Chart speed = 0.5 mm/s 



n 




25 

12.5 




,000 

500 



■r^nr 1 ^T', vn '""Pf ' wwr-fy^-f^ f *t 



L 



Mull 



-i I i_ 



20 



\- E 



p 10 - 



) 100 200 300 400 500 600 70 



300 



400 
TIME, s 



500 



600 



700 



800 



Figure 17.-Typical strip-chart data, 2.3 st/min (top) and 1 to 8 st/min (bottom). 



26 




Figure 1 8.-Coal fines buildup on conveyor deck. 



Throughout the test, the chain carried coal fines back 
underneath the conveyor decks. This material was ejected 
from the system at the point where the conveyor drive 
sprockets engaged the conveyor chain. It was only very 
fine material that piled up beneath the vehicles. It should 
be noted that no significant material loss was found around 
the hoppers at vehicle transfer points, even on those vehi- 
cles that were skewed at a steep angle. 

After this testing was completed, the amount of coal 
lost from each vehicle was weighed. Typical coal carry- 
back loss can be seen in figure 19. Table 8 shows the 
amount of coal lost under each vehicle and also quantifies 
that loss as a percentage of the total amount of coal con- 
veyed. As a percent of the total 682.1 st conveyed, loss 
from individual vehicles ranged from 0.026 pet in interme- 
diate vehicle 8 to 0.217 pet in intermediate vehicle 3. 
Average carryback loss was 0.063 pet per vehicle or 0.75 
pet total. Vehicle 3 tended to lose more material than 
the others and was therefore manually cleaned out during 
the test. This in itself increased the amount of loss by 
allowing more room for material to drop from the convey- 
or. It should be noted that because material became finer 
as the test progressed, rate of carryback loss increased. In 

Table 8.-Conveyor carryback loss 



Vehicle 



Loss, 
lb 



Portion of total 
conveyed, 1 pet 



Lead 

1 

2 

3 

4 

5 

6 

7 

8 

9 

10 

Discharge .... 
Total 

'Total conveyed, 682.1 St. 



614 
554 
747 

2,957 
755 
426 
475 
514 
355 
885 

1,241 
740 



0.045 
.041 
.055 
.217 
.055 
.031 
.035 
.038 
.026 
.065 
.091 
.054 



10,263 



.75 



an underground mine, coal will only pass through the 
system once. The percentage of fines will be much lower; 
therefore, carryback losses will be less. 

The conveyor gearbox temperatures of the discharge 
vehicle and intermediate vehicle 10 were monitored 
throughout the test program with Type K thermocouples. 
A temperature stabilization curve for gearbox lubricant is 
shown in figure 20. Data were obtained on March 28, 
1984, at haulage rate of st/min for a 135-min period. A 
stabilization curve at various loading rates could not be 
obtained because the system would not operate for a 
sufficient period of time to stabilize oil temperature. 

The temperature at which the lubricant temperature 
stabilizes gives an indication of the load on the gearbox 
and its condition. Temperature in intermediate vehicle 10 
stabilized at 131° F, with an ambient temperature of 56° F 
(or 77° F above ambient). This temperature is well within 
bounds of reasonable operating temperature. 

The discharge vehicle gearbox temperature did not 
stabilize during this period of time, but the slope of the 
curve was decreasing rapidly toward the end of the run. 
Temperature at the end of the run was 102° F above 
ambient. Under these conditions, discharge conveyor 
motor current and power consumption were 12 A and 7.8 
kW (10.5 hp), respectively, at 0-st/min haulage rate. 

The results of the March 14 through April 12, 1984 
conveyor tests follow: 

o A total of 682.1 st of coal was conveyed during 560.6 
min of operation. 

o Coal fines, generated by recirculation of the coal, 
were getting wetter on each recirculation in the closed- 
loop system because of the water sprays. The wetted coal 
fines increased the operating loads significantly, which 
caused a high failure rate of conveyor drive components. 
Because of the large number of mechanical failures and 
electrical problems, no more than 136 min of operating 
time was achieved in any single day (28 pet availability). 



27 



o Conveying horsepower requirements ranged from 5 
hp per vehicle at no load (0 st/min) to 18.4 hp per vehicle 
at 5 st/min (while conveying wet coal fines). 

o Specific energy required to transport the coal 
through the entire 228.8-ft system was 0.41 kW • h/st. 

o Almost no coal was lost through the system at vehi- 
cle transfer points. The only loss was through conveyor 
carryback, which represented about 0.75 pet of the total 
amount of coal conveyed. 

o Conveyor chain speed was 270.9 ft/min at 4-st/min 
haulage rate. Transport speed of discrete particles was 
234 ft/min at 4 st/min. 

o Carryback on the conveyor return deck for each 
vehicle was between 0.5 and 1 in. Some material was 
consolidated and cemented to the deck, some was loose 
(fig. 21). 



o Very hard, consolidated material had built up in the 
pelican beak in front of the conveyor drive shaft (fig. 22). 
This area had been previously cleaned using a slate bar 
through the cleanout ports. This method of cleaning the 
pelican beak is ineffective, as the material becomes hard 
packed and will not flow out the cleanout ports. Even 
when loose, material will assume a natural angle of repose, 
building up to the edge of the cleanout ports before any 
coal exits the area. This material eventually is carried into 
the hopper. 

o The amount of coal in the hoppers varied from 
vehicle to vehicle. Hard-packed coal fines were found 
throughout the hopper, especially between the chain guides 
and sides of the hopper, of intermediate vehicles 3 and 4. 
It was evident that this probably created a bind in the 
system, increasing the operating loads. 

o There were not a great deal of consolidated coal 
fines in intermediate vehicle 8. 



o Coal fines were tightly cemented into place between 
the chain guides and sideplates on the return deck of some 
vehicles. Material was so hard that the chain rode over 
the top of the coal. 



o It was found in intermediate vehicle 3 that the chain 
guides were spaced too close together and the chain links, 
rather than the chain rollers, were riding on the chain 
guides. This also increased operating loads. 




Figure 19.-Typical coal carryback. 



28 




• Discharge vehicle 
conveyor gearbox 

A Intermediate vehicle 
conveyor gearbox 



30 60 90 120 150 180 

TIME FROM INITIAL STARTUP, min 

Figure 20.-Conveyor drive gearbox temperature rise. 



July 31 to August 3, 1984 
Objective 

During previous haulage testing, it was found that one 
of the primary problems encountered was that of coal 
carryback. The coal carried back into the return deck 
increased power consumption and caused the system to 
periodically plug and shear drive pins, or cause other 
component failures. 

In an attempt to diminish the problems caused by 
carryback, it was decided to cut cleanout holes in the 
return deck to allow the coal carryback to exit, preventing 
its buildup. Therefore, cleanout ports were cut in the 
following areas of intermediate vehicles 3, 4, and 8 and the 
discharge vehicle: 

1. Hopper- -Coal that is carried back into this area can 
cause the conveyor to stall, therefore two 4- by 8-in holes 
were flame cut just behind the foot shaft. 

2. Pelican beak-Heavy buildup always occurs in this 
area. Even when the cleanout ports are removed, material 
will not clean itself from this area. Two 7.5- by 8-in 



cleanout ports were cut in the bottom of the pelican beak 
(fig. 23). 

3. Return cfec/c.-Material builds up on this deck, in- 
creasing chain flight friction, therefore, the built-in outby 
cleanout port covers and middle cleanout ports were re- 
moved. The expectation was that the material that would 
otherwise be carried back would drop out onto the next 
outby vehicle hopper. 

4. Area between chain guides and sideboard-Slots, 
approximately 0.75 by 6 in, were burned in the return deck 
between the chain guides and sideboards to allow material 
carried back by the chain links to exit. This area has a 
tendency to become plugged to a greater extent than the 
return deck. 

Coal buildup on the top decks was another problem 
that was addressed. Chain holddowns were fabricated 
from 2- by 2- by 0.25-in, 10-ft-long angle iron and bolted 
to the hopper sideboards of intermediate vehicles 3, 4, and 
8, and the discharge vehicle. 

The chain guides on the return deck of intermediate 
vehicle 3 were removed, allowing the chain flights to con- 
tact the return deck. It was expected that this would help 
purge the return deck of coal. 

Procedure 

The MUCH system was trammed into a circle with the 
discharge vehicle dumping onto the lead vehicle, the bridge 
conveyor was not utilized. The belt scale was unavailable, 
as it was assigned to another project. Otherwise, instru- 
mentation that was used in April haulage test was again 
utilized. Total conveyor system power, discharge vehicle 
conveyor power, discharge vehicle conveyor current, and 
system voltage were measured on the Gould model 481 
strip-chart recorder. 

The conveyor system was loaded with a fine mix of 
coal of an unknown size consistency from the Bureau's 
Hydraulic Transport Research Facility (HTRF) stockpile. 
Haulage rate was maintained at an estimated 3 to 6 
st/min. 

Results and Discussion 

The system was operated for several hours during 
July 31 to August 3, 1984, at haulage rates between and 
6 st/min (estimated). Because of the variability of the 
data and lack of haulage rate data, it was difficult to draw 
any definitive conclusions, but the system suffered no 
mechanical component failures (except shearpins) during 
this test period. 



29 




Figure 21 -Conveyor return coal buildup. 




# 





Figure 22.-Coal fines buildup in pelican beak. 



30 




Figure 23.-Cleanout ports In pelican beak. 



Generally, the steady-state power consumption was 85 
to 95 kW, discharge vehicle current was 12 to 13 A, and 
power was 8 to 8.5 kW at an estimated (rough) 3 st/min. 
Comparing these figures to the regression equations devel- 
oped for discharge vehicle current (I d , = 16 A) and total 
system power (P„ 123 kW) during the April 1984 haulage 
testing, it appears that the cleanout ports were effective 
to some degree judging by the reduced motor current and 
power consumption. Upon teardown of the upper decks, 
it was observed that there was much less coal left on the 
return decks and in the hopper return area than in the 
previous haulage testing in March. The regression used 
were 

Discharge vehicle current (I d ) = 12.17 + 1.26 Q, and 

Total system power (P,) = 62.8 + 20.2 Q, 

where Q = haulage rate, st/min. 



The problem with these cleanouts was that an excessive 
amount of material was lost through the cleanout holes 
(fig. 24). Appendix C shows the carryback loss for each 
vehicle. Carryback loss averaged 264 lb for the unmodi- 
fied (no cleanout ports) vehicles. The modified (cleanout 
ports added) vehicles averaged a 2,255-lb loss, or an 
unacceptable 8.5 times greater carryback loss; therefore, 
the cleanout ports were covered. 

Chain holddowns were effective in preventing coal 
buildup on the top decks. None of the modified vehicles 
had any appreciable buildup on the top decks. The con- 
figuration of the leading edge of the holddowns on the 
discharge vehicle had to be modified. The ramped shape 
of the leading edge caused coal to wedge between the 
holddown and chain, pinching the chain, increasing sliding 
friction, and stalling the conveyor. Cutting off this ramped 
transition area resolved the problem, returning the motor 
current to its normal level. Removal of the chain guides 
on the return deck of intermediate vehicle 3 was very 
effective in cleaning the deck. 



31 




Figure 24.-Coal carryback loss after haulage testing, July 31 to August 3, 1 984. 



September 10-11, 1984 

Objective 

Water sprays were fabricated to test the concept of 
spraying water in the return deck to minimize the problem 
of coal carryback plugging the system. Four spray bars or 
manifolds were fabricated (fig. 25) using eight Whirl jet 
3/8 BD 8 size 1 nozzles manufactured by Spraying Systems 
Co. Two nozzles were used per spray bar, 16 in apart. 
Two spray bars were installed in intermediate vehicle 8 
and two in the discharge vehicle. One spray manifold in 
each vehicle was inserted in the pelican beak through the 
cleanout ports. Nozzles were faced up and outby approxi- 
mately a 45° angle (fig. 26). The middle cleanout port in 
the return deck of both vehicles was removed and the 
manifolds were fastened into place with the nozzles facing 
up and outby. Cleanout ports used during the July 31 
through August 3 haulage test were covered. 

Procedure 

The MUCH system was placed in a circle and instru- 
mented as before in July and August conveying tests. 
Testing was conducted as before, adding coal as needed to 



achieve an estimated haulage rate of between 3 and 
6 st/min. 

Results and Discussion 

Water pressure at the machine inlet was 135 psi when 
the eight water sprays were functioning. Total flow rate 
measured was 2.69 gal/min or 0.336 gal/min per nozzle. 

The test did not run long before the coal was extremely 
wet, to the point of being a slurry. The material leaked 
out of the conveyors because of its liquid nature and the 
floor became extremely wet. This amount of water lost 
would present a serious problem underground, therefore 
the water spray manifolds were removed. 

Umbrella Miner Cutting Trials Support 

Objective 

To gain more information on the haulage characteristics 
of the MUCH system conveyors, it was decided to utilize 
the system to remove coalcrete cuttings that were pro- 
duced during umbrella miner cutting trials in the Bureau's 
Miner-Bolter Test Structure (MBTS). 



32 




Figure 25. -Water spray nozzles. 




Figure 26.-Water spray manifold in pelican beak. 



33 



Procedure 

The MUCH system was positioned to receive cuttings 
from the umbrella miner in the MBTS, to convey the 
material through the simulated workings, and then to 
discharge the cuttings from the bridge conveyor into the 
bucket of a front-end loader at the opposite end of the 
building. The system was instrumented to monitor the 
electrical system voltage, the total system electrical power 
in kilowatts, the discharge vehicle conveyor power in 
kilowatts, and the total system current requirements in 
amperes. Before testing, the conveyor system on each 
vehicle was run to verify proper operation and to remove 
any old coal fines from the conveyors. Each conveyor 
chain was oiled to free up any "frozen" links, and a new 
conveyor drive clutch was installed in the discharge vehi- 
cle. During this conveyor trial, coalcrete cuttings would 
be conveyed only once over the system, as opposed to the 
previous closed-loop type of haulage tests, which tended to 
overload the conveyors with a large quantity of fines from 
material degradation over a period of time. This trial 
would closely simulate actual coal haulage in an under- 
ground production situation. 

Results and Discussion 

Although the MUCH conveyor system performed al- 
most flawlessly during the trial and the only downtime was 
10 min because of one conveyor drive shearpin failure, the 
trial did not produce much useful information. Mechanical 
difficulties with the umbrella miner while trying to cut the 
coalcrete were the main problems. An insufficient amount 
of coalcrete material was loaded to tax the MUCH con- 
veyor system much above its power requirements while 
running empty. 

The maximum power requirement observed during the 
trial, except for startup power, was 25 kW (33.5 hp) for 
the total system power. When the system was running 
empty with no load, 18.5 kW (25 hp) was required. The 
maximum observed loading was only a 35-pct increase over 
the no-load conditions. The maximum load required to 
start the system was 40 kW (54 hp). 

During the previous closed-loop conveyor trials con- 
ducted in 1984, no-load (0 st/min) power consumption 
data were taken for the system; these did not include the 
bridge conveyor. Because of previous coal haulage trials, 
a large amount of coal fines was in the return deck and 
pelican beak areas of the vehicles, as well as wet fines that 
had accumulated on the conveyor decks. No-load power 
consumption data during those trials ranged from 45 kW 
to 68 kW (60 to 90 hp), which is approximately three times 
higher than what was observed in this test. Obviously, the 
accumulation of coal fines and conveyor chain lubrication 
have a significant influence on the amount of power re- 
quired to drive the conveyor system and the subsequent 
loads. 



Conveyor System Problem Summary 

Upon reviewing the results of the four tests, it became 
obvious that there was a major problem with the conveyor 
system ability to handle fines. The coal fines, built up on 
the conveyor top deck and on the return deck, ultimately 
overloaded the system to the point of failure. The addi- 
tion of water only made the problem worse. Cleanout 
ports in the return deck helped reduce fines buildup but 
lead to excessive loss of material from the system by 
dumping fines on the bottom. The closed-loop test setup 
was also contributing to the fines handling problems. 
Conveying coal in a closed loop continuously degrades the 
coal and yields all fines, which is not reflective of a nor- 
mal ROM product that the system would be handling at 
an underground production face. Based on these obser- 
vations a number of changes were made to the system as 
follows. 

o A conveyor chain slack adjustment mechanism was 
added to each vehicle to provide a simple means of keep- 
ing the conveyor chains tight so as to limit fines buildup 
on the decks. This modification is described in the 
"System Modifications" section. 

o Autoguard torque clutches were removed from the 
conveyor drive systems and replaced by Lovejoy couplings. 

o The conveyor drive shearpins were removed and the 
shearpin couplings were welded solid. Conveyor drive 
motor thermal overloads were resized to adequately pro- 
tect the drive components. 

o The conveyor chain holddowns were removed, 
redesigned, and reinstalled. 

Approximately 12 st of 3- by 1-in gravel was purchased 
for further conveyor system evaluation after the comple- 
tion of these modifications. Approximately 5 st of the 3- 
by 1-in gravel was dumped onto the system. The system 
was started and ran a very short period of time before it 
shut down on electrical overload trippings because a 
conveyor jammed. 

The jam was caused when the conveyor chain rode over 
a rock (fig. 27) and the rock was caught between the tail 
sprocket shaft and the chain flight (fig. 28). The primary 
cause was the large gap between the end of the conveyor 
deck and the tail sprocket shaft as shown in figure 29. 
This gap permitted material to drop between the end of 
the conveyor deck and the tail sprocket shaft. Material 
that was too large to fall through would either be broken 
by the next flight and exit the gap, stay in the gap momen- 
tarily and lift the conveyor chain as it passed over (which 
allowed material to get caught under other chain flights), 
or it would become jammed between the chain flight and 
the tail sprocket shaft. Fines that were small enough to 



34 



fall through the gap would mostly fall into the hopper of 
the next vehicle but a fraction would be carried by the 
return chain flights into the return deck to cause further 
jamming problems. 

The solution to these jamming problems was to (1) de- 
sign and install a plate on the conveyor deck to bridge the 
gap, and (2) redesign the conveyor chain holddowns on 
each vehicle. These modifications are fully described in 
the "System Modifications" section of this report and are 
shown in figures 30 and 31. 

Coal Conveyor Acceptance Test-October 29, 1986 

Upon the completion of the conveyor modifications on 
all vehicles, except the bridge conveyor, the system was 
run empty for a couple of hours to check operation and 
to break the system in before attempting to load the 



conveyors. After the initial break-in of the conveyors, the 
system was moved outside to a level open area and was 
trammed into a closed loop for testing. The bridge con- 
veyor was included in the loop. Approximately 8 st of 1- 
by 3-in gravel was loaded onto the conveyors and the 
conveyors were run in a closed loop for 30 min with no 
problems occurring. The rock was then conveyed off the 
system and 6 st of ROM coal was loaded onto the system 
for the next day's acceptance test. 

Objective 

The goal of the conveyor acceptance test was to con- 
tinuously convey coal at a rate of 8 st/min over an 8-h 
period while maintaining a system availability of 95 pet. 
For an 8-h test (480 min) this allowed only 24 min of 
downtime. 




Figure 27.-Conveyor chain riding over rock. 



35 




Figure 28.-Rock caught between shaft and flight. 





Figure 29.-Gap between deck and shaft. 



36 




Figure 30.-Conveyor deck plate. 




Figure 31 .-Conveyor chain holddowns. 



37 



Procedure 



DRAWBAR PULL TEST 



The test was conducted on October 12, 1986. The log 
of the test event is given in appendix D. The test was 
initiated with 6 st of ROM coal on the system. During the 
first 30 m in of the test an additional 6 st of coal was added 
to make up for the loss of fines as the coal degraded and 
to maintain 8 to 10 st of material on the system. A belt 
scale was not utilized during the test to avoid unnecessary 
downtime the additional belt might cause. 

Approximately 1 h into the test an additional 2 st of 
coal was added to the system. At this point in the test a 
total of 14 st of ROM coal had been utilized, approximate- 
ly 12 st was on the system and about 2 st had been lost as 
fines. As the test continued the coal kept degrading and 
fines were being lost, so after about 2 h of running time, 
3 st of the 1- by 3-in gravel was added and mixed with the 
coal. 

Because the test was being conducted outside, and to 
avoid problems previously encountered because of excess 
water, very little water was added to the material during 
the test. The water sprays were used only when the dust 
became excessive. As the test continued, the conveyor 
drive motors thermal overloads kicked out because of 
excessive loads, especially on the lead vehicle. The 
haulage rate was estimated between 12 to 13 st/min of 
mixed coal and rock. Approximately 2 st of material was 
removed from the system and the test continued. The test 
was terminated at 2:24 p.m. with over 8 st of material 
remaining on the system. 

Results and Discussion 

The test time was from 8:00 a.m. to 2:45 p.m., a total 
of 405 min. The total haulage time was 337 min and the 
total downtime was 68 min. The system availability for the 
test was 83.2 pet. The majority of the downtime, 58 min 
or 85 pet, was due to thermal overload trips on the con- 
veyor drive motors on the lead, intermediate vehicle 2, 
and intermediate vehicle 5. After the test the conveyor 
drive motor thermal overload setting on each vehicle was 
checked. On the three problem vehicles the settings were 
found to be lower than on the other vehicles, which con- 
tributed to the excessive tripping problem. 

The total amount of coal-rock mixture conveyed is 
estimated by 337 min of haulage time at an approximate 
rate of 10 st/min or 3,370 st. A total of 17 st of material 
was loaded on the system of which 2 st was removed and 
8 st remained on the system at the end of the test. There- 
fore, 7 st of material was lost during testing, mostly as 
fines, with some of the loss due to spillage at dumping 
points between vehicles. The total amount of lost mate- 
rial, 7 st, is equivalent to 0.2 pet of the total 3,370 st 
conveyed. 



Objective 

A drawbar pull test was conducted to measure the trac- 
tive effort produced by the MUCH system. The amount 
of tractive effort available gives an indication of the 
capability of the system to tram up grades and, if needed, 
to tow or pull a disabled or buried piece of equipment. It 
should be noted that the MUCH system is not intended or 
designed to be a towing vehicle and the use as such could 
lead to system damage. 

Procedure 

The drawbar pull of the MUCH system (without bridge 
conveyor) was measured by pulling against a 50-st-capacity 
Dillon dynamometer, which was anchored to a 35-st 
mobile crane (fig. 32). The dynamometer was attached to 
the discharge vehicle at the conveyor bridge mounting 
holes in the frame using a heavy chain. The dynamometer 
has a resolution of ±500 lb and a factory-stated accuracy 
of 0.5 pet of full scale. The brakes of the crane were 
applied during the pull test to keep the crane from mov- 
ing. The test was conducted outdoors on a heavily com- 
pacted, level, dry, dirt surface. The system was trammed 
and positioned into as straight a line as possible prior to 
being connected by chains to the dynamometer and crane. 

Results and Discussion 

The first pull test was conducted with no material in 
the conveyors. The empty weight of the MUCH system is 
approximately 117,800 lb. The tractive effort measured by 
the dynamometer was between 27,000 and 30,000 lb of 
steady pull. It was observed during this trial that some of 
the tires were spinning against the ground and on some of 
the intermediate vehicles the wheel rims were spinning 
inside the tires. This may be one disadvantage of having 
foam-filled tires as opposed to air-filled tires, the adhesion 
of the foam-filled tire to the rim is less in some cases than 
the friction between the tire and the ground, which can 
reduce tractive effort. 

The second pull test (fig. 33) was conducted after ap- 
proximately 10 st of material was loaded onto the MUCH 
system, which increased the total weight to approximately 
137,800 lb. The steady drawbar pull increased to about 
35,000 lb. After a few moments of tractive effort at this 
level three of the tram motor overloads tripped, shutting 
tram motors down. 



38 




Figure 32.-Crane and dynamometer used during drawbar pull test 




ROUGH TERRAIN TEST 



Objective 



Figure 33.-Drawbar pull test 



A rough terrain test was conducted to evaluate the 
performance of the system while tramming over areas 
with difficult bottom conditions. 

Procedure 

The test was conducted outdoors in an area with a 
surface of mixed mud and coal that had been saturated 
with water. A 2-ft-high by 13-ft-wide by 13-ft-long pile of 
mixed dirt, coal, rock, and 4-in by 4-in by 4-ft wood crib 
blocks was constructed as an obstacle. A roughly 10-ft- 
diam, 6-in deep water-mud hole was also part of the test 
course. The MUCH system was trammed repeatedly over 
the test course. 

Results and Discussion 

The system trammed over and through the pile of 
material and water-mud hole with no problems and no 
vehicle-to-vehicle binding or interference was observed. 
The front hopper on the lead vehicle acted as a plow to 



39 



help remove high obstacles prior to driving over them with 
the system. The wet, muddy, slippery bottom conditions 
had no adverse effect on the tracking ability of the system. 
As tramming over the test course continued, 6- to 8-in- 
deep ruts developed in one area that were of such length 
that two vehicles had their drawbars and undercarriages 
dragging through the mud. There was no noticeable loss 
of tram speed or tracking ability in this area. 

CANOPY LOAD TEST 

The operator canopy, which protects the lead vehicle 
operator, was redesigned and load tested. The canopy 
redesign and installation is described in the "System Mod- 
ifications" section of this report. 

As stated in the Code of Federal Regulations (30 CFR 
75.1710-1), a cab or canopy must have a "minimum 
structural capacity to support elastically: 

(1) A dead weight load of 18,000 pounds, or 

(2) 15 p.s.i. distributed uniformly over the plan view 
area of the structure, whichever is lesser." 

The elastic load criterion required by Federal law is 
based upon a combined statistical, analytical, and experi- 
mental investigation where it was found that a cab or 
canopy designed to meet this elastic load criterion has at 
least enough potential energy (in the form of available 
strain energy) to withstand the majority of roof falls as 
determined by the statistical analysis of all fatal roof falls 
from 1966 to 1972. 

The new MUCH system canopy was tested following 
MSHA guidelines. 7 The following items were used during 
the canopy load test: 

1. A Sensotec, Inc., 20,000-lb-capacity compression 
force transducer. 

2. A Fluke digital multimeter. 

3. A Honeywell Accudata 218 bridge amplifier. 

4. A 20-st-capacity hydraulic jack. 

5. A displacement dial indicator accurate to 0.001 in. 

6. A 12-in I-beam, 24 in long. 

7. The roof cart of the miner-bolter test structure 
(MBTS). 

8. Wood cribbing. 



Sawyer, S. G., and D. K. Brogam. A Testing Procedure for the 
Certification of Underground Protective Cabs and Canopies. MSHA 
IR 1002, 1974, 15 pp. 



Procedure 

In general, the test procedures, as outlined in MSHA 
IR 1002, were followed. Specifically, the lead vehicle on 
which the canopy is mounted was positioned directly under 
the roof cart of the MBTS and wood cribbing was used to 
support the canopy base and raise the lead vehicle's wheels 
off the floor. The top of the canopy was marked to show 
the middle ninth area of the total plan view area and the 
centroid of the canopy top was marked on the bottom sur- 
face of the canopy. A 12-in I-beam was cut and machined 
to cover and distribute the load over the middle ninth area 
of the canopy. The Sensotec 20,000-lb-capacity compres- 
sion force transducer, which was previously calibrated, was 
placed on the I-beam directly above the centroid of the 
canopy and connected to the Fluke digital multimeter and 
the bridge amplifier. The 20-st-capacity hydraulic jack was 
placed on the force transducer and also made to contact 
the bottom of the MBTS roof cart. The displacement dial 
indicator was mounted on the floor of the operators sta- 
tion and in contact with the centroid location on the 
underside of the canopy. 

The testing was conducted by utilizing the hydraulic 
jack to load the canopy in 1,000-lb increments, by reacting 
against the MBTS roof cart and by using the dial indicator 
to measure the canopy deflection. Once the canopy was 
loaded to 18,000 lb the load was removed and the residual 
deformation of the canopy was measured. 

Results and Discussion 

The results of the load test are shown graphically in 
figure 34. The maximum canopy deflection at a load of 
18,000 lb was 0.152 in at the centroid and the residual 
deformation at the centroid, once the load was removed, 
was 0.009 in. This residual deformation is 5.9 pet of the 
total deflection and falls well within the maximum allow- 
able residual deformation of 10 pet as stated in MSHA 
IR 1002. Therefore, the canopy is classified as substantial 
and is certifiable by a site-registered engineer. 

GROUND PRESSURE EVALUATION 

Objective 

A ground pressure evaluation was conducted to deter- 
mine the amount of pressure (pounds per square inch) 
that will be applied to the mine floor by the MUCH 
system tires during normal underground operations. This 
value is important in that the higher the pressure is at the 
tire-floor interface, the more likely the floor will be to 
deteriorate and ruts will be generated. 

Results 

For the evaluation, an average tire "footprint" was 
determined to be 10 in long by 8.5 in wide or 85 in 2 . Each 
vehicle has four tires for a total ground contact area of 



40 



340 in 2 . The unloaded weight of the lead vehicle was 
11,600 lb, the discharge vehicle weight was 13,200 lb, the 
intermediate vehicles weighed 9,300 lb each, and the 
bridge conveyor weight was about 3,000 lb. The ground 
pressure per tire for each vehicle empty and loaded with 
1 st of coal per vehicle is given in table 9. The ground 
pressure is the highest at the discharge vehicle with the 
bridge conveyor attached. 

Table 9.-Tire ground pressure for loaded 
and unloaded vehicles 

Vehicle Weight, lb Rear tire ground 

pressure, psi 

Lead: 

Unloaded 11,600 34 

Loaded 13,600 40 

Intermediate: 

Unloaded 9,300 27 

Loaded 11,300 33 

Discharge: 

Unloaded 13,200 39 

Loaded 15,200 45 

Discharge and bridge: 1 

Unloaded 14,700 48 

Loaded 17,700 54 

*Front tire ground pressures were 39 psi (unloaded) and 45 psi 
(loaded). 



SURFACE TEST SUMMARY 

The MUCH system was evaluated and modified during 
the surface test program to prepare the system for an in- 
mine production trial. The following are highlights of the 
surface test program. 

o The 12-unit MUCH system successfully trammed 
inby and outby through a simulated underground work 
area with 20-ft-wide entries and crosscuts at 90° to each 
other and 48-ft-square pillars. During these tracking and 
retracking trials, the system tended to drift toward the 
inside radius of the 90° turns. When tramming in the inby 
direction the maximum observed variance of any vehicle 
from the path of the lead vehicle was 68 in. When 
tramming in the outby direction the maximum observed 
variance of any vehicle from the path of the discharge 
vehicle was 101 in. The system tracked and retracked with 
enough consistency to operate in an underground mine of 
similar dimensions. 

o A maximum tracking variance of 35 in was observed 
during the simulated production cycle trial when the 
system was operating behind a simulated continuous miner 
making a 20-ft cut, a lift change, and second 20-ft cut. 



o The time to complete a simulated place change was 
3 min 20 s while tramming over a distance of 245 ft. 




KEY 

Load, Displacement, 
I0 3 lb in 






0.000 


1 


.007 


2 


.017 


3 


.027 


4 


.036 


5 


.044 


6 


.053 


7 


.062 


8 


.072 


9 


.076 


10 


.086 


1 1 


.095 


12 


.103 


13 


.120 


14 


.123 


15 


.127 


16 


.136 


17 


.142 


18 


.152 



■ Max deflection = 0.152 in 

• Residual deflection = 0.009 in 



0.04 0.08 0.12 0.16 

DEFLECTION, in 



Figure 34.-Lead vehicle operator canopy load test, load 
versus deflection. 



o A noise level survey was conducted. At the lead 
vehicle operator compartment, a noise level of 102 dBA 
was measured when the conveyor system was operating 
with no coal. At a haulage rate of 3.7 st/min of coal, 
98 dBA was measured; at a haulage rate of 7.4 st/min, 
97 dBA was measured at the operator station. The noise 
levels measured indicate that some form of hearing pro- 
tection will be required for the MUCH system operators 
to maintain their 8-h noise exposure to 90 dBA or below. 

o The conveyor chain speed was measured on a num- 
ber of vehicles. The average conveyor chain speed with no 
load was 273 ft/min. The chain speed at a haulage rate of 
7 st/min was 271 ft/min. The average startup time for the 
conveyor system was 1.36 s at no load and 1.86 s at a 
4-st/min load rate. 

o After a number of conveyor tests and numerous 
modifications on conveyor system, an acceptance test was 
conducted. During the test, the conveyor system operated 
for 337 min out of a total test time of 405 min (83.2 pet 
availability), the haulage rate averaged 10 st/min of mixed 
coal and rock, and the total amount of material conveyed 
was approximately 3,370 st. 



41 



o A drawbar pull test was conducted in which the 
empty MUCH system exerted a steady pull of 30,000 lb 
tractive effort; when the system was loaded with 10 st of 
material, the drawbar pull increased to 35,000 lb. 

o During a rough terrain test the system trammed 
through a 2-ft-high mixture of mud, coal, rock, and crib 
blocks, a 6-in-deep water-mud hole, and numerous 6- to 8- 
in-deep ruts with no loss in tram speed or tracking ability. 

o The redesigned operator canopy was load tested 
with a static load of 18,000 lb applied to the center ninth 



area of the canopy. The maximum deflection was 0.152 in 
and the residual deformation once the load was removed 
was 0.009 in. The canopy is classified as substantial and is 
certifiable by a State registered engineer. 

o A ground pressure evaluation was conducted. The 
highest ground pressure was at the rear axle of the dis- 
charge vehicle with the bridge conveyor attached and a 
full load of coal on the discharge vehicle and bridge con- 
veyor. The maximum rear tire ground pressure was 54 psi. 



SYSTEM MODIFICATIONS 



Numerous modifications were made to the MUCH sys- 
tem to prepare it for in-mine trials and to correct observed 
deficiencies found during surface testing. A summary of 
the modifications is given in appendix F. The system 
modifications are discussed in the following sections. 

MSHA EXPERIMENTAL 
PERMIT APPROVAL 

In order to allow an in-mine trial of the MUCH system, 
an experimental permit had to be acquired from the Mine 
Safety and Health Administration (MSHA) Approval and 
Certification Center. The effort to acquire an experiment- 
al permit was initiated by a meeting of Bureau, Boeing 
Services, and MSHA personnel on January 24, 1985, at the 
MSHA Approval and Certification Center, Triadelphia, 
WV. The purpose of the meeting was to discuss the pro- 
cedures of permit acquisition, to find out what materials 
MSHA currently had concerning the MUCH system 
(JMMD had at one time initiated an approval effort), and 
to acquire information on any new changes in permit 
requirements that might effect the system. The formal 
application for the experimental permit was sent to MSHA 
under Company Application Code No. 112525 on July 29, 
1985. After numerous drawing changes, many MUCH 
system revisions, and much correspondence with MSHA, 
experimental permit approval was received for the system 
on October 1, 1986. Experimental permit No. EP-541-0 as 
issued for the MUCH system had a duration of 6 months 
and with a request to MSHA, was renewed on April 7, 
1987, for an additional 6-month period. 

During the approval process, the following electrical 
deficiencies were defined and resolved to meet MSHA's 
requirements. 

The electrical control circuits of the MUCH system, as 
designed by JMMD, operated at a voltage of 460 V ac 
line-to-line. As stated in a MSHA memorandum, dated 
October 25, 1982, "The voltage of alternating current 
control circuits shall not exceed nominal 120 V line-to- 
line." The 460-V ac MUCH control circuits were in 
violation of MSHA regulations and had to be dropped to 



120 V to be approvable. The electrical control circuit was 
modified by installing a Mining Control, Inc., 24610 460-V 
primary to 120- V secondary, 2-kV«A, stepdown control 
transformer contained in an approved (X/P 2504-1) enclo- 
sure mounted on the discharge vehicle adjacent to the 
system's main electrical control box. The low-voltage 
secondary was equipped with 15-A fuses to protect the 
control circuitry wiring. The control voltage reduction 
necessitated replacement of the main circuit breaker (CB- 
1), shunt trip coil, and all magnetic coils for the motor 
control contactors and time delay relays. 

The emergency stop circuit was not-fail safe because 
normally open contacts at the emergency stop switches had 
to be closed to activate the shunt trip coil and open the 
contacts of the main circuit breaker. If the emergency 
switch contacts were not functional because of corrosion or 
mechanical problems, the circuit could not be completed 
to activate the shunt coil and open the breaker contacts. 
The emergency stop circuit and shunt trip coil were 
changed so that a closed circuit was required to activate 
the shunt trip coil and hold the main circuit breaker 
contacts closed. The contacts of emergency stop switches 
were changed from normally open to normally closed so 
that when the switches were activated, the contacts would 
open and break the circuit to the shunt trip coil and open 
the main breaker contacts. 

The trailing cable of the MUCH system was originally 
a 1/0, three-conductor, 90° C cable rated at an approxi- 
mate ampacity of 219 A for underground use. The total 
full load motor current of the system is approximately 
374 A and a normal load that would be expected during 
underground operations over a given period of time would 
be approximately 250 A. Obviously, the 1/0 trailing cable 
was of insufficient ampacity to meet MSHA requirements. 
The 1/0 cable was replaced with a three-conductor, 3/0 G- 
GC, 2,000- V, 90° C cable with a maximum ampacity of 
294 A. The 1/0 size power cable between the discharge 
vehicle and intermediate vehicle 10 was replaced with 3/0 
cable; and 1/0 power cables between intermediate vehicles 
10 and 9 and between 9 and 8 were replaced with 2/0 
cable. 



42 



The electrical fault ground-check circuit was original- 
ly connected to grounding studs in the explosion-proof 
(XP) box of each vehicle. This was not an MSHA ap- 
proved circuit because an open ground fault circuit to an 
individual vehicle would not be detected if the ground- 
check circuit to any other vehicle was continuous. The 
circuit was revised such that the ground check could only 
be connected to the ground stud in the lead vehicle, 
thereby providing one continuous path instead of a parallel 
path at each vehicle. 

The headlight circuit on the discharge vehicle was 
originally energized through a connection box located on 
the right side of the vehicle, which did not meet MSHA 
specifications. The circuit was changed to include a two- 
pole lever action pushbutton switch mounted in an XP 
enclosure in lieu of the connection box. 

Inspection of the MUCH electrical system components 
revealed that all electrical motor thermal overload relay 
heater elements were improperly sized to adequately 
protect the motors during normal operation. All heater 
elements were replaced with properly specified elements. 

The following refurbishment, replacement, and/or 
repair actions were required prior to the MSHA on-site 
inspection. 

1. Electrical cable conduit and clamps were installed 
or relocated to provide sufficient protection to electrical 
cables. 

2. The routing and suspension of electrical cables 
between vehicles was modified to prevent mechanical 
damage. 

3. The trailing cable ground-check circuit was checked 
for proper operation. 

4. Certification tags on electrical components were 
replaced or relocated to permit easier access during field 
inspection. 

5. The cable for the pager phone system was inspected 
and repaired as necessary. 

6. All hose conduit damaged during the test program 
was replaced or repaired. 

7. All cable packing glands were checked for proper 
clearances. 

8. All control box covers and edges were polished and 
checked for proper clearances. 

9. All electrical components in control boxes were 
tagged for easy identification. 

10. All electrical connections and wire terminals were 
checked and tightened as necessary. 



11. Electrical layout and wiring diagrams were updated 
to include all electrical modifications and corrections. 

OPERATOR CANOPY REDESIGN 

The operator canopy, which protects the lead vehicle 
operator, was redesigned and load tested. During tram- 
ming operation of the system in the EMTA, the top of the 
canopy came into contact with the simulated ribs of the 
EMTA thereby bending the canopy's vertical supports 
(fig. 35). Although this original JMMD canopy design met 
MSHA requirements for vertical load capacity, it was very 
weak under horizontal loading conditions. After the three 
vertical supports were deformed by rib contact, the ability 
of the canopy to support the MSHA-required vertical load 
was reduced to an unacceptable level. This required rede- 
sign of the canopy to increase the strength and number of 
canopy supports. 

In the new design (figs. 36-37), the three vertical sup- 
ports were replaced by four vertical supports. The new 
supports are constructed of rectangular steel tubing instead 
of threaded rod as was used in the original design. Each 
of the four vertical supports consists of a length of 4-in- 
long by 2-in-wide by 1/4-in-thick rectangular steel tubing 
welded to the canopy and a length of 5-in-long by 3-in- 
wide by 3/8-in-thick rectangular steel tubing welded to the 
base of the operator compartment. The 4- by 2-in tubing 
fits inside the 5- by 3-in tubing and is secured in place by 
locking pins inserted through holes drilled in both lengths 
of tubing. The modified canopy has an operating height 
range of 42 in to 56 in, and is adjustable in 3.5-in incre- 
ments over the height range. The canopy top strength was 
increased by using 1/2-in-thick steel plate instead of the 
original 3/8-in plate and by increasing the size of the steel 
tubing frame. Additional structural modifications were 
made to increase the strength of the floor area. Upon the 
completion of the installation of the modified canopy the 
structure was load tested and met MSHA requirements. 

ELEVATION OF CONVEYOR DECK 

In the original JMMD design there was approximately 
1.5 in of clearance between the forward edge of a vehicle's 
coal receiving hopper and the bottom surface of the con- 
veyor deck of the next forward vehicle when the system 
was on level terrain, as shown in figures 38 (top) and 39 
(top). During tramming and maneuvering of the system it 
became obvious that additional clearance was needed in 
this area. 

When tramming over uneven terrain, especially when 
one vehicle is twisting along the longitudinal center line 
relative to the next as when one wheel runs over a crib 
block, the vehicles bind together in this throat area (fig. 
40, top) and the tracking ability of the vehicle is lost. To 
alleviate this problem, the frame of each vehicle was 
modified to raise the discharge end of the conveyor deck 
by approximately 5 in. This modification increased the 



43 



height of the throat area from 13 in (fig. 40, top) to 18 in 
(fig. 40, bottom) and increased the hopper clearance from 
1.5 in to 8 in (figs. 38-39, bottom). 

CABLE HANDLING TRAY 

Cable handling trays were fabricated and installed 
through the drawbars between vehicles. The trays (fig. 
41), fabricated from 8-in C-channel, pivot in the middle to 
help reduce cable and hose damage. The trays support the 
power cable, control cable, telephone cable, and water 
hose between vehicles. 

DISCHARGE VEHICLE STEERING SYSTEM 

The hydraulic steering unit on the discharge vehicle was 
inadequate to allow the system to tram through the EMTA 
in the outby direction. A greater steering angle was 
needed; therefore, the linkage system was redesigned and 
modified to increase the steering angle and still utilize the 
existing cylinder. A bell crank arrangement was manufac- 
tured, and the steering cylinder was relocated. Figure 42 
shows the arrangement. 



EMERGENCY SHUTDOWN SYSTEM 

During the surface tests it became obvious that an 
improved method of emergency shutdown of the MUCH 
electrical system was required. The system as received had 
only three locations of emergency shutoff, one at the lead 
vehicle operator station and one on each side of the dis- 
charge vehicle; these two locations are approximately 
225 ft apart. In an underground situation neither the lead 
nor discharge vehicle operator can see the middle third of 
the system. If a problem occurred at a location not visible 
to either operator there was no way to shut down the 
entire system. To remedy this situation during surface 
testing, an emergency shutdown pull cord was installed 
along the center of the system for the entire length, which 
allowed complete system shutdown from any location 
(fig. 43). It must be noted that although this shutdown 
system worked well during the surface testing of the 
MUCH system, it was not designed for underground use. 




Figure 35.-Original operator canopy. 



44 




Figure 36.-Lead vehicle operator canopy. 




Figure 37. -Operator in lead vehicle. 



45 







«^»* 








Figure 38.-Clearance between hopper and deck-original Figure 39-Conveyor deck. Before modification, 1 .5-in throat 

design (top) and modified design (bottom). clearance (top); after modification, 8-in throat clearance (bottom). 





r I \r\rtr A/ 




Figure 40.-Height of throat area-original design (top) and modified design (bottom). 



46 




Figure 41 -Cable handling tray. 




Figure 42.-Discharge vehicle hydraulic steering system arrangement 



47 




Figure 43.-Emergency shutdown pull cord. 



CONVEYOR CHAIN SLACK 
ADJUSTMENT MECHANISM 

During coal haulage trials, problems occurred from 
fines buildup on the conveyor decks, which significantly 
increased haulage power requirements. One major cause 
of these problems was the slack in conveyor chain. As the 
MUCH system was originally designed, the only way of 
removing slack from the conveyor chain was by physically 
removing one or more links from the chain; but this could 
only be accomplished when 2 in or more of slack was 
available. Under extreme circumstances 2 in of chain slack 
was sufficient to allow a buildup of approximately 5 in of 
fines on the conveyor deck. To correct this problem, a 
conveyor-chain slack adjustment mechanism was designed 
and installed on the vehicles. 

The mechanism, as shown in figures 44 and 45, consists 
of two adjustment screws-one on each of the conveyor 
sideboards. The sideboards, which support the tailshaft of 
the conveyor chain, were cut approximately 65 in from the 
tailshaft, and the sideboard mounting holes were slotted 
to allow 5 in of movement. The adjusting screws are 
mounted on the forward half of the sideboard, and used to 
position the rear half of the sideboards and tailshaft to 



adjust conveyor-chain tension. The sideboard mounting 
bolts hold the sideboards and tailshaft in position when 
the chain adjustment is completed. 

CONVEYOR CHAIN HOLDDOWNS 

In the course of evaluating the conveyor system it be- 
came obvious that an effective conveyor chain holddown 
system was needed to help prevent fines buildup on the 
conveyor decks. Prior to the final conveyor acceptance 
test, a chain holddown system was installed on all vehicles 
(fig. 31). The holddowns consisted of 2- by 2- by 0.25-in 
steel angles bolted to the conveyor sideboards with 0.25- 
in clearance between the top of the chain and the bottom 
of the angle. The holddowns extend the entire length of 
the conveyor deck from the end of the loading hopper to 
about 4 in past the tail sprocket, which minimizes the 
amount of material that could be pulled in between the 
chain and the holddowns. A 2.5- by 0.25-in steel strip 
was welded between the legs of the angle to eliminate 
another potential area of material buildup. The chain 
holddown mounting holes in the rear portion of the con- 
veyor sideboard were slotted to allow chain slack adjust- 
ment without removing the holddowns (fig. 46). 



48 




Figure 44. -Conveyor chain slack adjustment screw. 



IS 



O O QifcuufiQipi ifly it iu)n m 



( l"l C 



»»■■' '"■■■ ■ " ■■£ 



^Q 



"'Hill, 111 IN 





PLAN 



r-^5-^ 



Center line 



-"fl 1 - 



QEHOfiggl 



iJii'm-iji'i' i'B 



-w* 



igr- 



"#" 



ELEVATION 

Figure 45.-Conveyor chain slack adjustment mechanism. 



49 




Figure 46.-Chain holddown mounting hole. 




Figure 47.-Conveyor deck extension plate. 



50 



CONVEYOR DECK EXTENSION 

As discussed in the "Conveyor System Tests" section, 
there was a material conveying problem because of a gap 
between the end of the conveyor deck and the tail sprocket 
shafts (fig. 29). This gap allowed chunks of material to get 
caught and either jam the conveyor or elevate the chain. 
This problem was amplified by the addition of the chain 
tensioning system by which the chain tension was increased 
by moving the sideboards. Increasing the chain tension 
also moved the tail sprocket shaft away from the conveyor 



deck, increasing the size of the gap. This problem was 
corrected by the addition of a 0.25-in-thick steel plate to 
bridge the gap between the conveyor deck and the tail 
sprocket shaft (fig. 47). The plate lies on the conveyor 
deck with the rear edge as close as possible to the tail 
sprocket shaft. The two outside edges of the plate are 
welded to the two conveyor sideboards so that the posi- 
tioning between the plate and the sprocket shaft remains 
constant when the conveyor chain tension is adjusted. The 
leading edges of the plate are chamfered so as not to bind 
the conveyor chain. 



CONCLUSIONS 



Surface testing showed that the MUCH system has the 
potential to substantially increase the productivity of a 
room-and-pillar mining system. The following items are 
recommended to improve its general functioning in an 
underground operation. 

o An improved emergency shutdown system should be 
designed and installed on the system to allow electrical 
deactivation from any point along either side of the con- 
veyor train. 

o Voice-activated headsets should be used by the 
lead and discharge vehicles operators to permit hands-off 
communications. 

o State-of-the-art noise reduction technologies should 
be applied to reduce the conveyor system noise levels. 

o MSHA-approved in-line cable connectors should be 
installed in the power and control cables between vehicles. 
These connectors would reduce vehicle change out time. 

o All conveyor drive components should be strength- 
ened by at least 50 pet. 



o A positive means of locking the foam-filled tires to 
the rims is needed to prevent the tires from spinning on 
the rims. 

o A cable reel mounted along the panel belt for the 
MUCH system trailing cable would help to limit the possi- 
bility of the discharge vehicle tramming over the cable. 

o An adequate supply of spare parts should be avail- 
able at the mine site to minimize downtime. Appendix E 
is a recommended spare parts list. 

o A thorough operator's training program should be 
conducted for the face production personnel prior to the 
start of an in-mine trial. 

o Additional lighting along the length of the system 
would be an advantage for both tramming and safety. 

o Audio signal on the system to warn the persons 
when it moves forward or rearward. 

o Rearview mirror in the lead vehicle operator cab to 
permit the operator to see behind the lead vehicle without 
turning around. 



51 



APPENDIX A.-MUCH SYSTEM SPECIFICATIONS 

The MUCH system is designed to be used in an underground room-and-pillar mining system. Because it is a versatile 
system, it can be used in a highwall operation also. 

Total system length (12 units) ft . . 250 

Estimated total weight (12 units including bridge conveyor) lb . . 120,800 

Power, 3-phase, 60 Hz V ac . . 460 

Minimum turning radius ft . . 24 

Mine configuration: 

Entry-crosscut width ft . . 18-20 

Entry-crosscut angle deg . . 60-90 

Minimum working height, in: 

With canopy on lead car 60 

Without canopy on lead car 48 



Vehicle units 



Lead vehicle 



Intermediate 
vehicle 



Discharge 
vehicle 



Frame: 

Length ft 

Active length ft 

Height in 

Width ft 

Canopy (adjustable) in 

Conveyor: 

Chain speed ft/min 

Chain width in 

Trough height in 

Motor hp 

Capacity st/min 

Tram: 

Speed ft/min 

Motor hp 

Wheel size 

Tread width ft 

Brakes 

Drive 

Steering: 

Forward 

Reverse 

Communications 

Hydraulics: 

Pump 

Motor hp 

Headlights: 

Number 

Voltage V ac 

NAp Not applicable. 
'28.25 ft with bridge conveyor. 
*22 ft with bridge conveyor. 



23.25 




21.75 


'21.75 


19.75 




19 


2 19 


41 




44 


41 


6.5 




6 


6 


42-56 




NAp 


NAp 


280 




280 


280 


30 




30 


30 


9 




9 


9 


15 




15 


15 


12 




12 


12 


80 




80 


80 


7.5 




5 


7.5 


8.25 by 15 


8.25 


by 15 


8.25 by 15 


5 




5 


5 


Disk, hydraulic 


Spring activated, 
electrically 


Disk, hydraulic 




released 






Front wheel 


Front wheel 


Rear wheel 


Manual 


Automatic 


NAp 


Automatic 




NAp 


Manual 


Pager phone 


Optional on 1 


Pager phone 




unit 






1 unit 




NAp 


1 unit 


1 




NAp 


1 


2 




None 


2 


11 




NAp 


11 



52 



APPENDIX B.-CONVEYOR TEST BREAKDOWN AND REPAIR LOG 



Cumulative 
run time, 


Av loading 
rate, 


Machine 


mm 


st/min 




3/14/84: 
10 ... 




2 


MUCH 


18 ... 




2 


HFB 


50 ... 




2 


HFB 


3/15/84: 
59 ... 




2 


MUCH 


59.1 . 




2 
2 


MUCH 


63.1 . 




MUCH/HFB 


64.1 . 




1 


MUCH 


65.1 . 




1 


MUCH 



Description 



3/16/84: 
68.6 


1.5 
1.5 


MUCH 


69 


MUCH 


76 


1.6 


MUCH 


3/20/84: 79 





MUCH 


3/28/84: 215 







3/29/84: 
283 


2 


MUCH 


284 


2 


MUCH 



Breaker on intermediate vehicle 8 tripped out, reason 
unknown. Breaker was reset. 

Main breaker on HFB tripped out. Unable to reset 
breaker. Problem was diagnosed as a poor 
connection on source side of the breaker. 
Connections were tightened, resolving the 
problem. 

Main breaker on HFB tripped out, reason unknown 
Breaker was reset. 

Clutch slipped on intermediate vehicle 10. Vehicle 
was jogged back and forth to reset the clutch and 
unjam the conveyor. 

Intermediate vehicle 9 breaker tripped, reason 
unknown. Breaker was reset. 

Power center tripped out, reason unknown. Breaker 
was reset. 

Intermediate vehicles 2 through 10 and the discharge 
vehicle breakers tripped out, reason unknown. 
Intermediate vehicle 9 sheared a pin. Shearpin 
was replaced and the breakers were reset. 

Key sheared on conveyor drive shaft on intermediate 
vehicle 9. Keyseat was damaged and driveplate 
bore was galled. To repair the conveyor 
driveshaft, the driveplate was welded solidly to 
the driveshaft. 

Clutch slipped on intermediate vehicle 10 and was 
reset. 

Clutch slipped on intermediate vehicle 10 and was 
reset. 

Breakers on intermediate vehicles 7 and 8 tripped 
out, reason unknown. Intermediate vehicle 9 
sheared a pin. Breakers were reset and shearpin 
was replaced. 

Discharge vehicle speed sensor shaft seized and failed 
the stub shaft on end of tailshaft (speed sensor 
sprocket drive). Speed sensor was electrically 
bypassed. Replacement speed sensor and tailshaft 
were later replaced. 

No breakdowns. 

Breaker on intermediate vehicle 4 tripped out, reason 
unknown. Breaker was reset. 

Breaker on intermediate vehicle 8 tripped out, reason 
unknown. Breaker was reset. 



See explanatory notes at end of tabulation. 



53 



Cumulative 

run time, 

min 



Av loading 

rate, 

st/min 



Machine 



Description 



3/29/84: 
298 .. 



3/30/84: 
299 .. 

341 .. 

341.5 . 

346.5 . 

360.5 . 
362.5 . 



4/03/84: 
364 .. 

367 .. 

370 .. 

374 .. 

375 .. 

4/04/84: 
375.2 . 



382.2 



HFB 



MUCH 
HFB 

HFB 

MUCH 

MUCH 
MUCH 



3 


MUCH 


3 


MUCH 


3 


MUCH 


3 


MUCH 


3 


MUCH 



MUCH 



MUCH 



Main breaker on HFB tripped out. Problem was 
a poor HFB connection on load side of the 
breaker. Breaker was overheated in that area. 
Connection was tightened, resolving the tripping 
problem. 

Breakers on intermediate vehicles 6 and 7 tripped 
out, reason unknown. Breakers were reset. 

HFB conveyor hydraulic motor hose failed. The 
crimp failed, allowing hose to blow out end of 
the fitting. Crimp-type staple-lock fitting was 
replaced. 

HFB was jammed. Wet coal overloaded conveyor 
system. Hopper was emptied manually. Also, 
1 hydraulic fitting on conveyor motor was leaking 
its O-ring was replaced. 

Breaker on intermediate vehicle 1 tripped out, reason 
unknown. Discharge vehicle sheared a pin during 
system troubleshooting. Breaker was reset and 
shearpin was replaced. 

Breaker on intermediate vehicle 4 tripped out, reason 
unknown. Breaker was reset. 

Intermediate vehicle 3 conveyor drive chain broke 
and taper lock bushing bolt (on conveyor drive 
sprocket) sheared. Motor fan was also hitting 
against guard. Chain was repaired; bolts and fan 
were replaced. 

System shut down for unknown reason. System 
started up without a problem. 

Intermediate vehicles 7 and 8 breakers tripped out, 
reason unknown. Breakers were reset. 

Intermediate vehicles 4 and 6 breakers tripped out, 
reason unknown. Breakers were reset. 

Intermediate vehicles 4 and 6 breakers tripped out, 
reason unknown. Breakers were reset. 

Intermediate vehicles 4, 6, and 7 breakers tripped 
out, reason unknown. Breakers were reset. 

Intermediate vehicles 1, 2, 3, and lead vehicle did 
not start. Problem was traced to a loose wire 
in electrical XP box on intermediate vehicle 4 
(control power circuit). Instantaneous overload 
in breaker was set to its highest level, eliminating 
the nuisance tripping problem in intermediate 
vehicle 4. 

Breaker on intermediate vehicle 3 tripped out, reason 
unknown. Breaker was reset. 



See explanatory notes at end of tabulation. 



54 



Cumulative 

run time, 

min 


Av loading 

rate, 

st/min 


Machine 


4/04/84: 
402.2 


1.5 
1.5 
1.5 

1 

.8 
.8 
.8 
3 

3 

3 
3 

4 

4 

5 
5 


MUCH 


422.2 


MUCH 


436.2 


MUCH 


4/05/84: 
439.2 


MUCH 


443.2 


MUCH 


446.2 


MUCH 


447.2 


MUCH 


517.2 


MUCH 


04/06/84: 
518.2 


HFB 


525.2 


MUCH 


525.7 


HFB 


531.7 


HFB/MUCH 
MUCH 


536.7 


4/09/84: 
537.2 


MUCH 


548.2 


MUCH 



550.2 



MUCH 



Description 



Intermediate vehicle 3 sheared a pin. Pin was 
replaced. 

Clutch on intermediate vehicle 1 slipped. It was 
reset. 

Clutch on intermediate vehicle 1 slipped; it could 
not be reset. Clutch was taken out and replaced 
with a solid coupling (Lovejoy). 

Intermediate vehicle 6 sheared a pin. Shearpin was 
replaced. 

Clutch on intermediate vehicle 10 slipped. It was 
reset. 

Breaker on intermediate vheicle 1 tripped out, reason 
unknown. Breaker was reset. 

Breaker on intermediate vehicle 1 tripped out, reason 
unknown. Breaker was reset. 

Breakers on intermediate vehicle 5 and the discharge 
vehicle tripped out (thermal overload). After 
discharge vehicle motor cooled, breakers were 
successfully reset. 

HFB was jammed. Ran conveyor back and forth 
to clear hopper. 

Intermediate vehicle 1 stopped. Reason unknown. 

HFB was jammed. Manually shoveled the system 
and ran conveyor back and forth to clear system. 

HFB was jammed. Manually shoveled the system 
and ran conveyor back and forth to clear system. 
Intermediate vehicle 1 also stopped, reason 
unknown. 

Clutch on intermediate vehicle 10 slipped; it would 
not reset. Clutch was removed and replaced with 
a solid coupling (Lovejoy). 

Intermediate vehicle 3 stalled. It appeared that the 
overload trip was caused by fines buildup. 

Breaker on intermediate vehicle 8 tripped out, reason 
unknown. Breaker was reset. Intermediate 
vehicle 10 conveyor drive gearbox overheated. 
Temperature after 15 min of operation was 239° F. 
Haulage rate was 5 st/min average. It was found 
that the fines buildup around the conveyor drive 
chain had caused it to seize, significantly 
increasing the system load, and thereby increasing 
the gearbox temperature. After end of run on 
April 9, chain was removed, cleaned, and 
reinstalled. Gearbox no longer overheated. 

Breaker on intermediate vehicle 8 tripped out, reason 
unknown. Breaker was reset. 



See explanatory notes at end of tabulation. 



55 



Cumulative 
run time, 

min 

4/09/84: 
550.7 

551.7 

552.2 

4/12/84: 
553.2 

554.2 

556.2 

556.3 

556.4 

556.5 

556.6 

560.6 



Av loading 

rate, 

st/min 



Machine 



Description 



5 


MUCH 


6 


MUCH 


6 


HFB 



5 


MUCH 


4 


HFB 


4 


MUCH 


4 


MUCH 


4 


MUCH 


4 


MUCH 


4 


MUCH 


4 


MUCH 



Breaker on intermediate vehicle 8 tripped out, reason 
unknown. Breaker was reset. 

Breaker on intermediate vehicle 8 tripped out, reason 
unknown. Breaker was reset. 

HFB conveyor hydraulic motor hose blew out of 
crimp fitting. (This hose was on the opposite port 
to hose that failed on 3/30/84.) Hose was 
replaced. 

Lead vehicle sheared pin. Shearpin bushing was also 
damaged. Pin and bushing were replaced. 

HFB was jammed. Ran conveyor back and forth 
to clear system. 

Breakers on intermediate vehicles 7 and 8 tripped 
out, reason unknown. Breakers were reset. 

Breakers on intermediate vehicles 7 and 8 tripped 
out, reason unknown. Breakers were reset. 

Breaker on intermediate vehicle 7 tripped out, reason 
unknown. Breaker was reset. 

Breaker on intermediate vehicle 8 tripped out, reason 
unknown. Breaker was reset. 

Breakers on intermediate vehicles 7 and 8 tripped 
out, reason unknown. Breakers were reset. 

Breakers on discharge vehicle and intermediate 
vehicle 10 tripped out. Wire on load side of main 
breaker was loose. Connections were tightened 
and breakers were reset. 



HFB Hopper-feeder-bolter. 

MUCH Multiple-unit continuous haulage system. 

XP Explosion-proof. 



56 

APPENDIX C.-MUCH SYSTEM COAL CARRYBACK LOSS-CONVEYOR TEST- 
JULY 31 TO AUGUST 3, 1984 

Weight of coal, lb 

Lead vehicle: Outby end 369 

Intermediate vehicle 1: Conveyor drive shaft area 215 

Intermediate vehicle 2: Conveyor drive shaft area 318 

Intermediate vehicle 3 (modified), 2,504 lb lost: 

Hopper cleanout port 225 

Conveyor drive shaft area and pelican beak 681 

Midship cleanout port 780 

Outby cleanout port and intermediate vehicle 4 hopper cleanout port 818 

Intermediate vehicle 4 (modified), 1,853 lb lost: 

Conveyor drive shaft area and pelican beak 353 

Midship cleanout port and chain slots 668 

Rear chain slots and outby cleanout port 832 

Intermediate vehicle 5: Conveyor drive shaft area 155 

Intermediate vehicle 6: Conveyor drive shaft area (caps open) 147 

Intermediate vehicle 7: Conveyor drive shaft area 617 

Intermediate vehicle 8 (modified), 3,240 lb lost: 

Hopper cleanout port 253 

Conveyor drive shaft area 535 

Midship cleanout port, chain slot, and outby 2,632 

Intermediate vehicle 9 (vehicle misaligned) 109 

Average loss of unmodified vehicles 264 

Average loss of modified vehicles 2,258 

Increase in coal spillage 854 



57 



APPENDIX D.-CONVEYOR ACCEPTANCE TEST LOG-OCTOBER 29, 1986 



Clock Run time, min Delay time, min 

time Run Cumulative Delay Cumulative 

a.m.: 

8:00 

8:20 20 20 

8:23 3 23 

8:25 23 2 2 

8:27 2 25 2 

9:07 40 65 2 

9:08 65 1 3 

9:10 2 67 3 

9:37 27 94 3 

9:38 94 1 4 

9:45 7 101 4 

9:47 2 103 4 

9:50 3 106 4 

9:52 106 2 6 

9:55 3 109 6 

9:57 109 2 8 

10:00 3 112 8 

10:02 112 2 10 

10:05 3 115 10 

10:06 115 1 11 

10:16 10 125 11 

10:17 125 1 12 

10:26 9 134 12 

10:28 134 2 14 

10:30 134 2 16 

10:42 12 146 16 

10:45 146 3 19 

11:05 146 20 39 

p.m.: 

12:20 75 221 39 

12:30 221 10 49 

2:00 90 311 49 

2:15 311 15 64 

2:30 15 326 64 

2:34 326 4 68 

2:45 11 337 68 

NOTE.-Total test time-405 min, total run time-337 min, total down time-68 min. 



Event 



Start test, 6 st coal on system. 

3 st coal to system. 

System shutdown, lead vehicle 

conveyor motor. 
Restart system. 
3 st coal added to system. 
System shutdown. 
Restart system. 
2 st coal to system. 
System shutdown. 
Restart system. 
1 st rock to system. 
1 st rock to system. 
System shutdown, lead vehicle 

conveyor motor. 
Restart system, add 1 st rock. 
System shutdown, intermediate 

vehicle 2, conveyor motor. 
Restart system. 

System shutdown, intermediate 
vehicle 5 conveyor motor. 
Restart system. 
System shutdown, lead vehicle 

conveyor motor. 
Restart system. 
System shutdown, lead vehicle 

conveyor motor. 
Restart system. 
System shutdown, lead vehicle 

conveyor motor. 
Removed 2 st material from system. 
Restart system. 
System shutdown, lead vehicle 

conveyor motor. 
Adjusted thermal trip setting on 

lead vehicle. 
Restart system. 

Stop system, repair coupling on 

conveyor drive motor discharge 

vehicle. 
Restart system. 
System shutdown, lead vehicle 

conveyor motor. 
Restart system. 
System shutdown, intermediate 

vehicle 5 conveyor motor. 
Restart system. 
Stop test. 



58 

APPENDIX E.-RECOMMENDED SPARE PARTS FOR IN-MINE TRIAL 

Item Quantity 

7.5-hp tram motor, frame 215 TC 1 

5.0-hp tram motor, frame 215 TC 2 

Cone drive tram speed reducer 1 

1.0-hp hydraulic pump motor, frame 182 T 1 

Lovejoy couplings, 1-1/2-in x 5/8 in 6 

Speed switch chain, Morse 40 5B 80 pitch 4 

Speed switch sprocket 4 

Conveyor drive sprocket, 14-tooth 4 

15-hp conveyor motor, frame 254 TDZ 2 

Conveyor drive shaft with housing and bearings 2 

Conveyor toe shaft assembly 2 

Conveyor tail shaft assembly 2 

Tram brake assembly 1 

Headlight assembly 2 

Conveyor speed reducer, lead vehicle 1 

Conveyor speed reducer, intermediate and discharge vehicles 2 

Tires on rims 2 

Drive axle assembly (1 intermediate and 1 lead) 2 

Nondrive axle assembly 2 

Conveyor chain, including flights 2 



59 



APPENDIX F.-MUCH SYSTEM MODIFICATIONS SUMMARY 

Reason 



Modification 

Voltage of electrical control circuits were reduced from 
460 V ac to 120 V ac line-to-line. 



460- V primary to 120- V secondary, 2-kV»A stepdown 
control transformer installed in the electrical control 
circuit. 

15-A fuses added to control circuits. 



Main circuit breaker (CB-1) for system, shunt trip coil, 
and all magnetic coils for motor control contactors and 
time-delay relays were replaced. 

Emergency stop electrical circuit modified to provide fail- 
safe-operation. 

Size of system trailing cable increased from a 1/0, 3-con- 
ductor, 90° C rated cable to 3/0, 3-conductor, 90° C 
rated cable. 

1/0 size cable between the discharge and intermediate 
vehicle 10 replaced with 3/0 cable, and 1/0 cable be- 
tween intermediate vehicles 10 and 9 and beteen 9 and 
8 replaced with 2/0 size power cable. 

Electrical fault ground-check circuit was originally 
connected to grounding studs in the explosion-proof 
(XP) box of each vehicle. This was modified so that 
circuit was only connected to ground in the lead vehicle 
XP box. 

Electrical connection box for the discharge vehicle 
headlight circuit replaced with 2-pole lever action 
pushbutton switch mounted in XP enclosure. 

All electrical motor thermal overload relay heater elements 
were replaced. 

Lead vehicle operator canopy was redesigned and load 
tested. 

Frame structure of each vehicle was modified to elevate 
discharge end of conveyor deck by 5 in. 

Cable support trays were fabricated and installed through 
the drawbars between vehicles to support cables and 
water hose between vehicles. 

Discharge vehicle steering system was modified by 
fabricating a new bell crank arrangement and relocating 
steering cylinder. 



MSHA memorandum stated "The voltage of alternating 
current control circuits shall not exceed nominal 120 V 
line-to- line." 

Transformer was required to step down control circuit 
voltage from 460 V ac to 120 V ac. 

To protect the control circuit components from 
overcurrents. 

Changes required because of reduction of control circuit 
voltage from 460 V to 120 V. 



Safety. 



Amperage of 1/0 cable was insufficient to handle normal- 
ly expected amperage requirements of system. 



Do. 



Original grounding was not MSHA approvable because an 
open ground fault in an individual vehicle would not be 
detected. 



Connection box did not meet MSHA specifications. 



Original overload elements were improperly sized to 
adequately protect the motor. 

Original canopy was damaged during testing. 



To reduce vehicle-to-vehicle interference during tramming 
over uneven terrain. 

To help reduce cable and hose damage between vehicles. 



More steering angle was needed to improve steering ability 
of discharge vehicle. 



60 



Modification 

Pull-cord type emergency shutdown system was installed 
for surface testing of the MUCH. 

Conveyor chain slack adjustment system was designed and 
installed on vehicles. 

Conveyor chain holddowns were installed on vehicles. 

Conveyor deck extension plate was added to each vehicle 
to bridge gap between end of the conveyor deck and 
conveyor chain tail shaft. 

Shearpins located in conveyor drive mechanism were 
removed and shearpin couplings were welded solid. 

Autoguard torque clutches were removed from conveyor 
drive system and replaced with Lovejoy couplings. 



Reason 

To provide complete shutdown of entire system from 
either side along the length of the system. 

To reduce amount of fines buildup on conveyor decks. 



To prevent fines buildup on conveyor decks. 

To prevent material from getting jammed between end of 
conveyor deck and tail shaft. 



During conveyor testing, the shearpins proved to be 
unreliable. 

Clutches failed and could not be reset. 



* U.S. GOVERNMENT PRINTING OFFICE: 611-012/00,124 



INT.BU.OF MINES.PGH..PA 28977 



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